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By H. T. Kennedy, S. M. Avasthi

An equation developed for gaseous hydrocarbon mixtures predicts molal volumes with an average absolute deviation of 0.73 percent when applied to 264 natural gas and condensate systems including 2,043 PVT points. Another equation developed for liquid hydrocarbon mixtures predicts molal volumes with an average absolute deviation of 1.12 percent when applied to 346 crude oil systems including 1,759 PVT points. Both equations require composition of the mixture to be expressed as mole fraction of methane through heptanes-plus, hydrogen sulfide, nitrogen and carbon dioxide, together with the characteristics of the heptanes-plus fraction in addition to the temperature and pressure. The equations cover wide ranges of the variables involved, and their accuracy is considerably better than that of other available methods. The equations were differentiated to allow calculation of the coefficients of isothermal compressibility and isobaric thermal expansion. (In this paper the coefficient of isothermal com-pressibility and the coefficient of isobaric thermal expansion will be expressed as compressibility and thermal expansion coefficient, respectively.) Equations to calculate these quantities are presented. INTRODUCTION Calculations of reservoir performance for petroleum reservoirs require accurate knowledge of the volumetric behavior of hydrocarbon mixtures, both liquid and gaseous. Compressibilities are required in transient fluid flow problems, and thermal expansion coefficients are important in thermal methods of production. An accurate laboratory investigation of the PVT behavior of each reservoir fluid encountered would be costly and time consuming. For this reason various correlations for predicting fluid properties have been developed and recorded in recent literature. Correlations have been presented in the form of graphs, tables and equations. Since an increasing number of studies are being conducted with the aid of electronic computers, recent efforts have been directed toward development of correla-tions suitable for computer programming. Application of computers permits the use of more complex correlations which otherwise are not feasible. Moreover, methods for predicting reservoir performance, particularly those based on the compositional material balance, depend upon the capability of accurately expressing the molal volumes and other fluid properties as functions of pressure, temperature and composition. The coefficient of isothermal compressibility c is defined by

Jan 1, 1969

By E. B. Johnson

Cleveland-Cliffs lron Co., which operates Marquette lron Mining Co.'s Republic mine, conducted a test whereby 35,000 tons of self-fluxed pellets were produced in the grate-kiln system from specular hematite flotation concentrates. A brief description of the pelletizing plant is presented along with pertinent operating details. Date is included which compare operating conditions and the chemical and physical properties of the green and fired pellets for the self-fluxing pelletizing test with the normal acid pelletizing operation. The success of self-fluxed sinter in enhancing the operation of the blast furnace during the past few years has reasonably led to the assumption that the use of self-fluxed pellets might result in even greater furnace efficiencies. A great share of development work in agglomeration has been directed towards techniques related to the production of self-fluxed pellets, but only a limited tonnage of self-fluxed pellets had been produced on a commercial scale until this past year when The Cleveland-Cliffs Iron Co. produced 35,000 tons of such pellets by the grate kiln system at the Republic mine during the period May 21-June 5, 1962. REPUBLIC'S GRATE-KILN PLANT The grate-kiln pelletizing process consists of four stages, namely: feed or balling circuit, traveling grate, rotary kiln and annular cooler (Fig. 1). The balling section is conventional and utilizes four 9x30-ft balling drums in closed circuit with 5x14-ft vibrating screens having 3/8-in. rod decks. The traveling grate is 12 ft wide by 126 ft long and has 14 windboxes — six in the drying section and eight in the preheat section. The rotary kiln is 15 ft ID shell and 114 ft long. The annular cooler is 40 ft in diameter and has 12 windboxes. The cooler is divided into two zones by a moveable curtain wall. The high temperature air from the first zone is used as secondary combustion air in the kiln. Normally the only heat supplied is through the burning of No. 6 oil in the single kiln burner. Hot gases pass counter to the pellet flow through the system. The exit gases from the kiln pass down draft through the preheat section through cyclones and are drawn through the drying bed prior to exiting via the stack. The temperatures are normally held within the following ranges: Kiln 2400-2445 °F Preheat 1850-1950°F Drying 480-520°F Exit gas 200°F Pellets 120°F The design capacity of the Republic mine grate-kiln system was 2500 Itpd. Sustained production in excess of 35% over design capacity has been achieved. No screening of the cooler product is practiced. PELLETIZING TEST CONDITIONS When a decision was made to proceed with a test at the Republic mine pelletizing plant to produce 35,000 tons of self-fluxed pellets, several potential failure factors had to be carefully considered. Research work on a batch and pilot plant scale indicated that the grate-kiln system at Republic would require some major changes in order to obtain optimum conditions. In many cases compromises were agreed upon as major equipment changes and revisions were not feasible. The pelletizing conditions which laboratory development tests revealed were required in order to insure successful production of self-fluxed pellets were: 1) Good mixing of limestone and concentrate and proper moisture control; 2) Higher gas flow rates in the preheat section than are normally used; 3) Higher preheat temperatures than normal; 4) Maximum kiln temperature 2350°F. There was just cause for concern about balling problems that might occur. Approximately 10% of dry pulverized limestone was to be added to the specular hematite concentrate just ahead of the Pekay mixers on the belt feeding the balling drum. Large scale

Jan 1, 1963

By W. J. Verner, J. R. Lucas

Billions of tons of bulk solid materials are processed through our industrial plants each year, and the tonnage is steadily rising. It has been estimated that for every dollar spent in industry as a whole, approxinlately 30 cents is consumed in the handling of bulk solid materials. It is in this category that tremendous savings are possible through the use of improved techniques. Since World War 11, the mining, steel, chemical, and grain industries, among others have made use of increased materials-handling automation in order to gain some of these savings. However, in a chain of automatic operations, every link must be completely reliable, for if there is a single failure, the whole system immediately becomes ineffective. Only too often the weak link in the chain proves to be the flow of bulk solids from bins. To contribute to the fundamental knowledge in this area, a research study has been continuing at The Ohio State University. The principal objective of the investigation has been to study the internal stresses existing in confined bulk solids. LITERATURE SURVEY Literature studies have revealed that many outstanding people have made contributions to the field of bulk solids. For instance, C. F. Jenkin,1 suggested: a. The pressure on a bin wall is not directly proportional to the depth beneath the surface of material in the bin. b. The vertical pressure distribution on the base of a bin is not actually uniform but is a variable. c. Arches form in bins as a result of very small deformations of the material in the bin. This deformation causes a series of gaps to appear, and consequently causes the disappearance of diagonal forces that cannot act across these gaps. The removal of the diagonal forces causes horizontal forces to be introduced which seem to be the chief cause of arching. 11. A. Janssen,4 in 1895 in connection with his experiments on bins, derived a differential equation to describe both the lateral and vertical pressure in bins with vertical sides. This represents a significant contribution to bulk solid theory, and thus its solution is reproduced as follows: L = KV = RW/U' [l-e -U'KY/R] I where L is the lateral pressure of the bulk solid in psf., V is the vertical pressure of the bulk solid in psf., K is the ratio of the lateral pressure to the vertical pressure at any point, R is the hydraulic radius of the bin in feet, W is the density of the bulk solid in lbs. per cubic feet, U' is the coefficient of friction between the stored material and the bin wall, and Y is the head of material above the point in question in feet. In 1903, Jamieson,6 in an elaborate series of experiments on full sized bins measured vertical and lateral pressures by hydraulic diaphragms. He determined many values for mediums of corn, wheat, peas, and flaxseed. In 1951, Caughey, Tooles, and scheer3 measured both lateral and vertical pressure in deep bins by using strain guages on steel bands. Values were calculated for cement, gravel, sand, soy beans, shelled corn

Jan 1, 1961

By D. Turnbull, H. N. Treaftis

Precipitation of tin from Pb-Sn alloys (lead-rich) occurs by the nucleation and growth of hemispherical cells which consist of tin lomelloe interspersed in the depleted solid solution. Nucleation and growth of these cells and the tin interlamellar spacing at various compositions and temperatures have been investigated. The interlamellar spacing increases approximately as the reciprocal of the logarithm of the supersaturation ratio. For constant initial tin concentration the concentration of cell nuclei increases with decreasing temperature to a limiting temperature-independent value. The kinetic results support the view that the cell growth rote is governed by the diffusion of tin along the cell boundary. In the main, the results confirm the authors' earlier deductions, drawn from resistometric and calorimetric data, about the precipitation mechonism. IT has been shown1, 2 that tin (ß) precipitates from lead-rich (a) PbSn solutions by the nucleation and growth of cells (cellular or discontinuous precipitation). These cells consist of tin lamellae, with a spacing 1, interspersed in a depleted a solution. Reorientation of a also accompanies cell growth. The kinetics of cellular precipitation of tin from lead have been measured by resistometric2, 3 and ca1orimetric1, 3 techniques. The time-transformation isotherms (t equals time; x, volume fraction of specimen transformed) of the initial rapid precipitation, which drains about 60 pet of the tin from solution, have been described by a kinetic law having the form x = 1 —exp (—bt n) [I] where n usually has the value 3.0. It was shown" that this law and the other kinetic facts can be explained satisfactorily if it is assumed that the cells grow at a constant rate and that all the cells nucleate in a time negligible in comparison with the period of cell growth before impingement. Thus, for hemispherical cells b = (2/3)pNG3 [21 where G is the rate of cell growth in cm per sec, and N is the number of cell nuclei per unit volume. The rapid rate of cell growth at low temperatures can be accounted for"' if it is assumed that the dissolved tin drains to the lamellae edges by diffusion along the sweeping cell boundary rather than by diffusion through the a crystal (volume diffusion). For cell growth governed by cell boundary diffusion it was shown, approximately G=2(X° - X° / X° ) DBA / l2 where A equals the thickness of the cell boundary; D R, the coefficient of diffusion of tin atoms within the boundary; and X° and X, are the atom fractions of tin in the supersaturated and saturated solutions, respectively. Zener8 had already shown that, where

Jan 1, 1959

By S. H. Ward, R. A. Barker

THE Juniper Prospect is in Carleton County, N. B., at approximately 46" 31' north latitude, 67" 20' west longitude. During the summer of 1955 an area in west-central New Brunswick was selected for aerial electromagnetic prospecting with American Metal Co.'s helicopter-mounted equipment. Generalized, small-scale government geologic maps had suggested that this area was located along the western border of the northeast-trending granitic backbone of northwestern New Brunswick and might be underlain by the Ordovician-Silurian sediments and volcanics that are the host rocks for the recently discovered sul-fide deposits of the Bathurst-Newcastle district. The helicopter electromagnetic equipment located an anomalous conductive area, coincident with a small aeromagnetic closure, about five miles west of the town of Juniper. Subsequently 30 mining claims (total area about 6600 x 8000 ft) were staked and company personnel conducted geological, geophysical, and geochemical investigations on the property. The claims and immediately contiguous areas were mapped geologically, and an area of about 2000 x 4000 ft was soil-sampled and also surveyed by ground electromagnetic and gravity methods. Finally six diamond drillholes, totaling 3500 ft in length, were drilled to test the anomalous conditions discovered by the several exploration techniques used. The discovery of base metal sulfides during drilling was a technical success for the exploration methods, but the quantity of sulfides was too small to be profitably exploited.

Jan 1, 1959

By W. R. Hibbard

Evidence is presented which confirms previous findings that models of solution strengthening depending solely on lattice parameter changes are incomplete. Direct evidence for the Suzuki interaction of chemical effects was obtained from alloys with constant lattice parameter, for which solution strengthening was nearly independent of temperature from 300° to 20°K. SOLUTION hardening principles have been reviewed recently by Parker and Hazlett.' In addition, the interactions of dislocations and solute atoms were reviewed by Cottrell.' It is not proposed to review this information again here, but to refer only to those details which are pertinent to the developments which have occurred since these papers were published. Literature Survey To set the stage, it has been known for some time that solid solutions are stronger than the pure parent metals, and that the strength increases with increasing amounts of alloying element. One of the most extensive earlier investigations was that of Lacy and Gensamer,3 who reported that the solid solution strengthening of iron is a power function of the composition, that the relative strengthening is a function of the solubility limit, and that the rate of strain hardening is a function of the amount of solution hardening. Thus, the slope of the plastic portion of the stress-strain curve increases with increasing yield strength. An approach to the development of principles resulted from work by Gulyeav,4 who reported that the strengthening effect of solid solution alloys is periodic in character when plotted as a function of the atomic number of the solute. A more quantitative approach was to relate the amount of solution hardening to the lattice parameter of a solid solution. This was done quite early by Norbury,5 and later by Brick, Martin, and Angier,6 Frye and Hume-Rothery,7 and French and Hibbard. This relationship is only approximate in character. The next development was to relate the amount of solution hardening to lattice stress, as proposed by Frye and co-workers.9 They calculated what they called ionic overlap from the lattice parameter changes represented as a change in volume, using Bridgman's hydrostatic compression equation. The resulting correlation was somewhat improved as compared to the simple correlation with lattice parameter. The next step was carried out by Dorn, Pietrow-sky, and Tietz,10 who suggested an effect of valence differences in addition to the effect of lattice parameter. Their proposed parameter shows flow stress as a single valued function of the concentration and a relationship involving these two atomic variables. In addition, they contributed new data on the temperature dependence of solution hardening which indicated that it was a strong function of temperature for aluminum alloys. Allen, Schofield, and Tate11 carried the valency effect even further by reporting data indicating that with a given solvent element the stress-strain curve is a function of electron density, entirely neglecting the size factor, which varied by a factor of two. This evidence led to a quantum mechanical calculation by Cottrell, Hunter, and Nabarro,12 based on the formation of a dipole around an edge disloca-

Jan 1, 1959

By C. H. Wayman, D. W. Fuerstenau

The effect of bubble size and concentration of certain reagents on the terminal velocity, shape, path, and drag coefficients of single air bubbles in distilled water has been investigated. Bubbles of a certain size range rise considerably faster than Reynolds number-drag coefficient relationships predict, whereas a small amount of frothing agent reduces their terminal velocity. The rapid velocity at which bubbles rise in distilled water appears to result from slip at the boundary and from circulation within the bubbles. Surface-active agents retard bubble motion through prevention of both circulation and slip at the boundary. A surface tension concentration gradient helps maintain bubble sphericity. HE gas phase is one of the indispensible in-gredients in flotation operations. Flotation depends on the collision of an air bubble and a mineral particle in a pulp and their ability to remain in contact long enough for adherence to take place. Before it is possible to understand the mechanics of air bubble-mineral particle encounter, it is necessary to learn more about the nature of bubbles themselves before collision with a mineral particle has taken place. To date, little work has been done on air bubbles in flotation systems. In 1945 Fahren-wald' presented a study of the role of frothers on air bubbles and on aeration in flotation, and recently Wark and Sutherland' discussed the work of Rosenberg- in relation to flotation. The research reported in this present article was undertaken to investigate systematically the effect of bubble size and the concentration of certain reagents on the terminal velocity, shape, path, and drag coefficient of single bubbles in distilled water. Most of the work is concerned with the effect of a-terpineol on bubble motion in water, but studies were made with potassium chloride, potassium hydroxide, potassium ethyl xanthate, and potassium amyl xanthate. Experimental Method and Materials: To obtain free bubble rise and reduce wall and surface effects, the investigators used a tank of 5½x5½x30-in. internal dimensions, filled with water to within 4 % in. of the top.' The tank was made of acrylic plastic so that the walls would have good optical properties. Distilled water, which was re-distilled in a block tin still and saturated with air before each test, was used in all the experiments. All solutions were made with reagent-grade chemicals. During each experiment the water was at room temperature, and during the course of an experiment the temperature did not fluctuate more than 0.5°C. Since the experiments were performed over a period of six months, the temperature varied between 25° and 27°C. The temperature at which each series of experiments was conducted is recorded with the data. The experimental apparatus is presented dia-grammatically in Fig. l. Individual air bubbles were generated from purified compressed air through a capillary of 0.003-cm internal diameter at a rate

Jan 1, 1959

Jan 1, 1966

By R. W. Stahl

Employees and dollars are necessary to all enterprises and any force, such as fire, which destroys either, can bring very serious consequences, including business failure. Since everyone acknowledges the necessity of conserving human life, we will devote most of our paper to the economic factors involved. A coal-mine fire is considerably more difficult to combat than a fire in a surface building because of the combustible surroundings in the mine, the explosive gases emitted from the coal, the suspension of explosive dust when mine timbers burn and the unsupported roof falls, the limited space in which approach to the fire can be made, and the limited avenues of escape. A survey of coal-mine-fire protection indicated that a rather low value was being placed on property and equipment. This prompted comparison with other industries. The comments made in this comparison are not intended as an indictment of the coal industry, since we do have mines that are well protected against fire, but the comments on conditions at the mines visited are factual and may help others to improve similar conditions. SURVEY PROCEDURES Plants of the steel, aluminum, rubber, chemical, and petroleum industries were surveyed, and all facets of the fire-protection scheme were covered. The industrial plant invariably had four times as much water readily available for instant use than the mine. In many mines where water is available, no equipment for applying it is provided, such as hose, nozzles, and adapters, so that time is lost in collecting this equipment. (See Table I.) EXTINGUISHERS Many more and generally larger portable extinpishers are available at the plant as compared to the mine. The extinguishers are generally placed nearer to sources of danger. In addition, respiratory protective equipment is always available at strategic points in plants but only in a few mines. (See Figs. 1 and Table 11) INSPECTING AND TESTING EQUIPMENT Plant water facilities are checked weekly and flushed at 6-month intervals, and hose is pressure-tested at least yearly; some mines check water yearly, but hose is not tested unless used. Fire extinguishers are inspected weekly in plants and completely overhauled yearly; one mine had monthly checks and overhauls only after use. FIRE BRIGADES AND TRAINING Every plant had designated fire-fighters and some complete full-time brigades; at one mine a chief mechanic was designated as being in charge of fires, and other depended entirely upon officials who might not be familiar with the equipment. COST OF FIRES AND PERCENTAGE OF CAPITAL INVESTMENT IN FIRE PROTECTION The direct cost of recent mine fires ranged from $128,000 to $240,000; indirect costs, such as lost tonnage, idled equipment, lost orders, and the like might run the cost to over $500,000. (See Figs. 2 and 3).

Jan 1, 1961

[Executive J. V. W. REYNDERS, Chairman WILLIAM H. BASSETT E. DEGOLYER I-I. C. BELLINGER RICHARD PETERS, JR. Finance GEORGE OTIS SMITH, Chairman H. A. GUESS F. JULIUS FOHS Admissions GEORGE D. BARRON, Chairman DAVID LEVINGER P. A. MOSMAN H. K. MASTERS J. A, SINGMASTER Alternates F. G. BREYER C. W. HAMILTON F. E. PIERCE]

Jan 1, 1928

By S. M. M. Safvi, A. Jowett

Small-scale continuous flotation tests are described in which the influence of 1) pulp density and 2) feed rate on the rate of flotation are investigated. The results provide further evidence of first-order kinetics for flotation and indicate that tests at different feed rates provide a particularly simple method of determining rate constants. Flotation kinetics are discussed in general terms. Over a period of many years numerous research workers have studied the rate of recovery of valuable materials from flotation pulps with a view to deter-mining the mechanism of the froth flotation process. Froth flo ation is fundamentally a collision process involviug at some stage collision and then mutual attachment between air bubbles and mineral particles. It is possible to draw analogies therefore between the flotation process involving mainly macroscopic particles and the ordinary processes of chemical reaction kinetics involving collisions between atoms, molecules and ions. By adopting this analogy we may postulate a general equation representing the kinetics of the flotation process: -dC/dt =K.CC ,Aa (1) The rate constant, K, will be a complex function involving reagent concentrations, particle and bubble sizes, induction times and no doubt flotation cell design. However, these are factors which can be kept constant in any series of experiments, permitting investigation of the exponents a and c, which determine the order of the reaction occurring in flotation. The various research investigations carried out so far have all aimed at evaluating a and c and in the majority of cases only since the concentration of air has usually been kept constant in a series of tests, so that a relationship of the following type has been investigated: -dC/dt = K1 Cc (2) The techniques of both batch testing and continuous testing have been used to investigate this relationship. The technique of continuous testing, in which steady-state conditions are established, permits direct investigation of the above equation. In the case of batch testing, C is decreasing continuously

Jan 1, 1961

By N. R. Morrow, C. C. Harris

The experimental points which describe capillary pressure curves are determined at apparent equilibria which are observed after bydrodynamic flow has ceased. For most systems, the time required to obtain equalization of pressure throughout the discontinuous part of a phase is prohibitive. To permit experimental points to be described as equilibria, a model of capillary behavior is proposed where mass transfer is restricted to bulk fluid flow. Model capillary pressure curves follow if the path described by such points is independent of the rate at which the saturation was changed to attain a capillary pressure point. Aemodified suction potential technique is used to study cyclic relationships between capillary pressure and moisture content for a porous mass. The time taken to complete an experiment was greatly reduced by using small samples. INTRODUCTION Capillary retention of liquid by porous materials has been investigated in the fields of hydrology,l soil science,2 oil reservoir engineering,3 chemical engineering," soil mechanics,5 textiles: paper making: and building materials In studies of the immiscible displacement of one fluid by another within a porous bed, drainage columnsl and suction potential techniques2 have been used to obtain relationships between pressure deficiency and saturation (Fig. 1). Except where there is no hysteresis of contact angle and the solid is of simple geometry, such as a tube of uniform cross section, there is hysteresis in the relationship between capillary pressure and saturation.2 The relationship which has received most attention is displacement of fluid from an initially saturated bed (Fig. 1, Curve R2), the final condition being an irreducible minimum fluid saturation Swr. Imbibition (Fig. 1, Curve A), further desaturation (Fig. 1, Curve R), and intermediate scanning curves have been studied to a lesser but increasing extent.2 ,3,9-12 This paper first considers the nature of the experimental points tracing the capillary pressure curves with respect to the modes and rates of mass transfer which are operative during the course of measurement. There are clear indications that the experimental points which describe these curves are obtained at apparent equilibria which are observed when viscous fluid flow has ceased; and any further changes in the fluid distribution are the result of much slower mass transfer processes, such as diffusion. Unless stated otherwise, this discussion applies to a stable packing of equal, smooth, hydrophilic spheres supported by a suction plate5 with water as the wetting phase and air as the nonwetting phase. The discussion can also be applied, in many

Jan 1, 1966

By R. F. Shurtz

The problem of estimating the precision of systematic samples from a mineral deposit is attacked by interpolating the quality, or other attribute measured, by using Fourier approximation. Such approximation using trigonometric functions yields a particularly simple expression for the variance of the mean value of the attribute sampled. The increasing volume of publication in which the science of statistics has been applied to the problem of predicting the quality of an ore deposit from sets of samples taken from it might lead one to believe that engineering judgment can be replaced by statistics. There is, in fact, an implicit application of the principle of indifference or of "no advance knowledge" which has been discredited in modern statistical theory and practice. These pages will show that it is not always advantageous, efficient, or safe to apply this principle. This problem will be dealt with by using series of orthonormal functions to describe the distribution of the quality of ore in space as contrasted with the more commonly used distribution of quality as a frequency function. Mining engineers are rarely well acquainted with the use of such series or of the functions comprising them. These uses are commonplace in the fields of communication and information theory and it is not surprising that applications taken from these fields should be found appropriate for the analysis of the information contained in a set of samples from an ore deposit. It is hoped that presentation of the following basic, but incomplete, results will generate wider interest in this method and a fuller development of it as well as of the necessary computational techniques. FUNDAMENTAL PREMISES Random variables in the strictest sense are not very common in mine sampling. The sets of samples are almost always taken systematically and are, hence, not subject to all the laws governing random events. Therefore, unless there is indeed random experimental error in the determinations of quality, the treatment of systematic samples of n members as though they were random samples of n members often leads to calculation of confidence limits for the mean quality which are larger than they should be with the result that unnecessary additional sampling is done. The reason for this will emerge in the following discussion leading to Eq. 24 and its generalization. In order to inquire into such questions on a rigorous basis and at the same time avoid arithmetic complications that would obscure the concepts involved, a one dimensional distribution of quality will be assumed. The quality, Q, in suitable units is expressed as a function of the distance, x, along some straight traverse of the deposit. Such a function

Jan 1, 1967

By Edwin Berry

THE consideration of the present value of a mine is of great importance when planning for the operation of a property,, and it is particularly so in the case of mines that have large ore reserves. At various times during the past 15 years, the writer has had occasion to make rather thorough studies of the subject, and has been greatly impressed by certain important points which these studies have brought out, points which, though they may be well known, are often lost track of in the rush to obtain low mining costs and high daily tonnage output. . It is the object of this paper to present briefly, and in a general way, some of these points, and especially those which have to do with the relationship which (1) ore reserves and plant capacity, (2) grade of ore mined, and (3) ore losses in mining, bear to the present value. Present value, sometimes called present worth or actuarial value, means the value today of the future earnings of an enterprise. Great industrial, manu-facturing or railroad enterprises, if properly handled, go on indefinitely and their value is directly dependent on their estimated annual earnings. Mines, or oil properties, on the other hand, have an earning capacity only until the ore or oil is exhausted, and their present value is dependent not only on their annual earnings but also on their life, which is the length of the period over which they can produce at a profit. A mine, or an oil property, is a "wasting asset," while an estab-lished manufacturing business is not, or should not be. In other words, the present value of a mining property. represents a sum of money on which there could be paid from operating profits a determined rate of interest, say 8 per cent., and in addition an amount which, if invested at compound interest as received, would at the end of the life of the mine pay back the present value. In his "Wealth and Income of the American People," Walter Renton Ingalls says: "The fundamental principle of mine valuation is the present worth of an annual dividend accruing during the period of years corresponding to the life of the mine."

Jan 7, 1922

By J. Gurland

A method of measuring the degree of contact between grains of 1 phase in multiphase alloys is derived. It is shown that the number and the areas of contacts between grains can be determined by metallographic analysis. Relations between the area of contact, the dihedral angle, and the mean free path of spherical grains are developed. It is often desirable to interpret the mechanical behavior of alloys in terms of structural features observed under the microscope. A quantitative description of microstructures of multi-phase alloys requires consideration of geometrical and topological factors such as the amount, size, and shape of grains and the continuity of the constituent phases. The number and dimensions of grains may be metallographically determined by applying the methods described by Smith and Guttman1 and by Fullman2 in their papers on the measurement of particle size and internal boundaries in opaque bodies. The present work intends to show that the same methods can be used to evaluate some of the parameters of phase continuity; namely, contiguity, number of contacts, and areas of contact of grains of 1 phase. Application of these concepts may be found in such problems as the calculation of electrical conductivity of 2-phase alloys, or the evaluation of the length of fracture path through grains of a brittle phase more or less dispersed in a matrix. I. DEFINITION The following definitions and derivations are based on 2-phase alloys, of which the units of the microstructure are grains. Contiguity is defined as the average fraction of surface area shared by 1 grain of a phase with all neighboring grains of the same phase. In a 2-phase alloy of constituents A and B, the contiguity of phase B is given by: where Cb is the contiguity of phase B aBß is the average area of contact between a grain of phase B and adjacent grains of phase B, aAB is the average area of contact between a grain of phase B and adjacent grains of phase A. The contiguity is a measure of the average degree of contact between grains of phase B in the 2-phase mixture. The ratio varies from 0 to 1 as the distribution of phase B changes from a completely dispersed to a fully agglomerated structure. Smith and Guttman1 have shown that the relative areas of 2 kinds of interfaces in a volume are directly proportional to the number of interfaces of each type intersected by a random line. The lines can be drawn on metallographically prepared surfaces as long as these are random sections through the sample. The contiguity ratio of phase B then becomes: Cb = 2NBB/NAB + 2N BB [2] where: NBB is the number of interfaces between grains of phase B intersected per unit length of random line. Nab is the number of interfaces between grains of phase B and grains of phase A intersected per unit length of random line. The factor 2 arises in Eq. [2] because each interface B-B is associated with 2 grains of B-phase, whereas the interface A-B involves only 1 grain of B-phase. The calculations of contiguity do not require any assumption as to volume fraction, particle size, or particle shape. Fullman2 gives an expression for the determination of the mean free path between particles. The mean free path PB between grains of phase A, i.e., through grains or agglomerates of phase B, may be found from the relationship: PB =2fB/NAB [3] where fB is the volume fraction of phase B. The calculation of the mean free path is independent of grain shape and grain size distribution. The calculation of the number of contacts and of contact area requires assumptions of grain shape, grain size, and contact shape. In the case of uni-

Jan 1, 1959

By O. E. Palasvirta

Several drum-type, low-intensity dry magnetic separators have recently been described 1,2,3,4 all of which share at least two generic similarities. The magnet consists of a multiplicity of poles of alternating polarity and, the speed of the drum shell, relatiw to the magnet, is such that the magnetics are agitated through rotation of the magnetic particles around their individual axes. The speed of the drum relative to the magnet (differential speed) determines the frequency of the field reversals. For effective agitation this frequency generally has to be above a specific value, which varies according to the mean particle size of the magnetics. With separators of small diameter, say less than 11/2 ft., a sufficient differential speed for treatment of fine magnetite can be attained only by rotating both the magnet and the drum separately; separators of large diameter may attain sufficiently high differential speeds with stationary magnets. Separators of large diameter thus have the advantage of simplicity of design. The capacity of a separator of this type is directly proportional to the speed of the shell. As the centrifugal force opposing the magnetic force increases at a slower rate than the peripheral speed of the shell when drum diameter is increased, large drums can be operated with higher peripheral speeds. Thus they have a higher capacity than smaller drums. Separation is also a statistical phenomenon, i.e. the grade of concentrate improves as the number of agitations or rotations of the magnetic particles per separation increases. As separators of large diameter have space for a large number of magnets they are best suited for production of optimum grade concentrate. The statements above are general clues For separator design. In the following an expression IS derived showing a more exact relationship between the pertinent separator variables. As both magnetic attraction and centrifugal forces are directly proportional to the volume of the affected bodies, complications in mathematical analysis, arising from variable particle size, are avoided by considering the force per unit volume of affected material. For the magnetic force per unit volume Runolinna (3) gives the following equation: V " 1 + N(ur, - 1) uo dr (1) where ur and uo stand for the relative permeability and the permeability of vacuum, respectively, N is the demagnetization factor, B the intensity of the magnetic field, and r the distance from the face of the magnetic pole. The permeability of a specific magnetite ore can generally be considered as a constant. For ground, liberated material N can also be treated as a constant, and the first simplification can be made by writing ur-1 1=k1. 1 + N(ur - 1) uD k1. Fig. 1 shows the field intensity curves for two U-magnets of different size, but of roughly similar geometry of cross-section as plotted against the distance from the plane of the pole faces. The curves

Jan 1, 1961

By J. B. Fletcher

The Miami mine first started operations in 1910. For convenience, the history of the orebody can be divided into the following categories (Fig. 1): 1) 1910 to 1925: 24.4 million tons of high grade ore, mined by top slicing, sublevel caving, etc. 2) 1926 to 1954: 102 million tons of low grade ore, mined by block caving.

Jan 1, 1961

By Francis Foley

INTRODUCTION DURING the summer of 1919, the late Dr. Henry M. Howe, then Chairman of the Division of Engineering of the National Research Council, organized a committee to obtain a better insight into the behavior of carbon steels under the influence of various treatments by a proper control of the time and temperature factors. During part of 1919 and 1920, work in connection with this investigation was carried on at Dr: Howe's laboratory at Bedford Hills, N. Y. During 1921 there was an enforced hiatus in the work, occasioned by the illness which resulted in Dr. Howe's death in 1922. At that time one of the present authors compiled the results of the work which had been done and this was published in 1923.1 Following the publication of this work, it seemed desirable to carry on in the same manner an investigation of steel of C 0.75 per cent., which steel had not been as fully treated as had the other steels of C 0.34 and 0.52 per cent., used in the published work. Furthermore, certain unusual results have been obtained in the treatment of C 0.52 per cent. steel by quenching and drawing, and repetition of this part of the work seemed advisable. This has been done during the past year at the School of Mines and Metallurgy of the University of Missouri at Rolla, Mo., by a cooperative undertaking of the School and the Bureau of Mines, working in conjunction with the Committee on Heat Treatment of Carbon Steels of the National Research Council. The present paper is a report of the results obtained in this work.

Jan 2, 1926

By B. A. Wilcox

ThERE are several reports in the literature which indicate that both solid-solution hydrogen and hydride precipitates can promote low-temperature em-brittlement of tantalum.1-3 For example, Imgram et al.2 showed that hydrogen additions in the range of 140 to 580 ppm promoted a ductile-to-brittle transition at slightly above room temperature in both wrought and recrystallized tantalum. This contrasts with the observation that tantalum of norma1 purity is very ductile, even at liquid-nitrogen temperature. Imgram et al. also found that hydrogen additions caused a deterioration of notch tensile properties of tantalum. The purpose of this communication is to present some limited evidence that hydrogen can also cause fatigue embrittlement of tantalum. The material used in this investigation was wrought, stress-relieved (1000°C, 1 hr), are-melted tantalum rod (5/8-in. diameter) containing the following impurities in weight percent: 0.005 C, 0.007 02, 0.0005 H2, 0.003 N2, 0.08 Cb, 0.01 W, and 0.01 Fe. Hydrogen charging was done in an atmosphere of flowing dry hydrogen for 3 hr at 600°C, followed by an homogenizing anneal in purified argon for 15 min at 600°C, and cooling in an argon atmosphere. Typical micro-structures of the hydrogen-free and the hydrogen-charged materials are shown in Fig. 1(a) and Fig. 1(b), respectively. Fig. 1(b) shows the appearance of hydride precipitates which, in many cases, formed preferentially parallel to the rod axis. The hydride dispersion was uniform across the diameter of the specimen. The fact that the precipitate was a tantalum hydride was tentatively confirmed by X-ray diffraction on a metallographically prepared surface. After hydrogen charging, chemical analyses by vacuum fusion revealed the hydrogen content to be 0.0295 wt pct. In addition, the oxygen and nitrogen contents were slightly raised to 0.0095 O2 and 0.005 N2.

Jan 1, 1964

By W. H. Bassett, C. H. Davis

In the commercial annealing of cartridge brass there are four points regarding which definite data are essential. They have to do with the correct interpretation of grain count in its relation to annealing temperature and, incidentally, to Brinell hardness. These points are: 1. The comparison of the grain sizes of two cartridge-brass mixtures: 69 copper, 31 zinc, 0.376-in. (9.5-mm.) gage; and 68 copper, 32 zinc, 0.130-in. (3.3-mm.) gage 2. The comparison of the grain sizes of annealed metal that had previously been reduced by rolling varying amounts; for instance, 20.0, 36.6, 50.9, and 59.1 per cent. 3. The determination of standards for grain sizes on annealed brass of the composition 68 per cent. copper, 32 per cent. zinc and 69 copper, 31 zinc. 4. The comparison of grain size with Brinell hardness on identical samples of annealed metal. In their comprehensive and thorough investigation of the recrystal-lization of cold-worked alpha brass on annealing,l Mathewson and Phillips have discussed the relations between temperature of anneal, degree of deformation, and structural alteration in alpha brass. They have also shown certain comparisons between the ordinary physical properties and the grain size of annealed brass. The purpose of the present investigation is mainly concerned with the grain size of cartridge brass, its relation to Brinell hardness, and the publication of sufficient data to enable those engaged in the inspection of such material to have a correct foundation upon which to work. The first alloy was taken from regular mill stock that had been rolled from 0.580 in. to 0.376 in. (14.7 mm. to 9.5 mm.) gage, a reduction of 35.1 per cent. This bar (No. 1) had the following composition: Copper, 69.20 per cent.; zinc, 30.76 per cent.; lead, 0.02 per cent.; iron, 0.02 per cent. The second alloy was also taken from mill stock of 0.325 in. (8.25 mm.) gage and was rolled as shown in Table 1, in order to get four bars reduced two, four, six, and eight B. & 5. numbers respectively. This second alloy

Jan 1, 1919