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By E. A. Reilly, G. R. Gardner, F. P. Lasseter

THE methods that may be applied to the treatment of slurries and water, as these are related to practical coal-preparation problems, are concerned essentially with the movements of solids suspended in water. It is assumed that there are no fundamental differences between the principles involved in the treatment of coal slurries and the phenomena encountered in general mineral-dressing practice, and the latter are presented in some detail. Basic principles ,are correlated with present practice for both the anthracite and the bituminous coal fields, and methods are discussed which are designed to utilize best all the material subjected to treatment and suggest a sound approach to the solution of problems relating to the return of clear water to process or to streams. In order to approach the practical phases of the subject with sufficient understanding of the basic principles, it is necessary to discuss two types of phenomena that are applicable. The first is flocculation and the second is sedimentation. Descriptive matter is given covering the various methods and types of equipment employed to conduct these operations, with emphasis on their application to water and coal-slurry treatment. DEFINITION OF TERMS Definitions of terms are consistent with modern developments, and on form, in general, with their usage in mineral-dressing practice. Because of variations in coal-preparation methods, it is impossible to define a rigid upper limit of particle size for a coal slurry or sludge which will apply generally. Limits vary somewhat with the particular treatment method employed at a given plant. For purposes of definition, this limit is chosen as the particle size for which preparation methods such as Chance cones and jigs are usually dispensed with, and other methods, such as the Hydrotator, tables, and froth flotation, are employed. Generally the limit will be 10 to 6 mesh. In assuming this variable upper limit of particle size in a coal slurry, it is recognized that the size ranges discussed will duplicate sizes dealt with

Jan 1, 1943

By Lyman Johnson

Single crystals qf copper in "single slip" orientatiorzs have been deformed in compression. During defortnation all of the independent deformation parameters have been measured. These parameters consist of thefive strain components and three components descrihing the lattice rotation. By a finite strain analysis these pararmeters , forrming a deformation gradient martrix, are related to the amounts of slip on each of the twelve slip systems. The results show that the amount of secondary slip is about equal to the amount of primary slip. This is an order of magnitude larger than has been believed previoutsly. ACCORDING to early theory and experiments, when a single crystal of a fcc metal is deformed in tension or compression it should deform by slip on only one slip system until the stress axis reaches a symmetrical orientation.' However. the observation of a large increase in the secondary dislocation density during ..single slip" makes it clear that some slip does occur on secondary systems. Knowledge of the amount and distribution of this secondary slip is essential to a complete understanding of the mechanisms of single-crystal deformation. Ahlers and Haasen 2 and Mitchell and Thornton1 have tried to detect the amount of secondary slip in single crystals of silver and copper, respectively. Each simultaneously measured the angle A, between the tensile axis and the primary slip direction and the length 1 of a gage section of the specimen after incremental amounts of deformation in tension. The measured A, was then compared with the theoretical single slip angle hp. given by sin Ap = j sin . hO where ?o was the initial angle between the tensile axis and the primary slip direction and lo was the initial gage length. In both sets of experiments a small but systematic difference between ?e and ?p was found. This difference must be due to the occurrence of secondary slip. However, as Mitchell and Thornton1 pointed out. nothing quantitative can be said about the amount and distribution of this secondary slip from the measurements that they made. The reason that no quantitative conclusions could be made is because no unique solution for the distribution of slip on the twelve fcc slip systems can be determined from only two measured deformation parameters such as A and 1. There are, in fact, eight independent macroscopic deformation parameters that can be measured when a single crystal undergoes a homogeneous deformation. Physically these can be thought of as the five finite strain components and the three angles describing the crystal lattice rotation. All eight of these parameters were measured by Taylor4,5 for aluminum deformed in tension and compression. At that time the concern was to show that slip occurs on {111 (110) systems in fcc metals, and the mathematics were not available to determine what slip distributions were compatible with the measurements. In this paper the mathematics6,7 are developed that allow the slip distribution to be determined from these measurable macroscopic deformation parameters. The analysis is applied to the measurements of the strain and lattice rotation of copper single crystals deformed in compression. The results show that the amount of secondary slip is an order of magnitude larger than had previously been thought. CRYSTALLOGRAPHIC DESCRIPTION OF A HOMOGENEOUS DEFORMATION The deformation of a solid body can be represented by a transformation matrix F that transforms the un-deformed state into the deformed state. Consider a vector X connecting two material points in the unde-formed material and the vector x connecting the same two material points after deformation, where both vectors are referred to the same set of Cartesian axes. The final vector x is related to the initial vector X by the equation: X = FS. [2] Eq. [2] can be considered as the equation defining F, which is called the deformation gradient matrix. Its components are: If the deformation is homogeneous, the transformation is linear and the components of F are constants. Using subscript notation, if P is the unit vector in the initial direction of a material line, the components of the unit vector p in the direction of the same material line after deformation are given by:

Jan 1, 1970

By J. J. Frawley, W. J. Childs

The dynamic nucleation of supercooled bismuth and Bi-Sn alloys has been studied over a frequency range of 15 to 20,000 cps. For low-frequency vibration, a minimum vibrational energy was required for enhancement of nucleation. Above this critical energy, the dynamic supercooling was less than static supercooling showing that vibration promoted nucleation. The amount of dynamic supercooling continued to decrease with increasing vibrational energy until a minimum or threshold value was reached. This minimum value of supercooling for nucleation remained constant joy all further increases in vibrational energy. For higher frequencies, similar results were observed. This behavior has been related to the necessity of cavitation for dynamic nucleation. When a liquid is cooled to a temperature below its equilibrium melting point, the solid phase is more thermodynamically stable. However, for solidification to occur, a two-step process, nucleation and subsequent growth of the solid phase, must occur. When a liquid is supercooled, that is cooled below the equilibrium melting point, the controlling process for solidification to begin is the rate of nucleation. Once nucleation has occurred, the solidification process is controlled by the rate of growth. Nucleation can be induced by two factors: either by a catalyst or by the use of mechanical shock. Numerous investigators1-4 have studied the effect of nucleation catalysis but much less systematic study has been made of nucleation by mechanical shock waves. The influence of vibrations on grain size in castings and ingots has been studied by many authors but no clear understanding of the mechanism or accurate prediction of the effect has been presented.5 It would be intuitively expected that the further the departure from equilibrium (i.e., the greater the supercooling), the easier it would be to induce nucleation. This has been quantitatively demonstrated both by walker6 and later by Stuhr,7 that the greater the degree of supercooling the easier it is to nucleate by a shock wave. Stuhr also attempted to obtain the mechanical energy required for nucleation of bismuth as a function of supercooling. He vibrated a crucible containing supercooled metal at low frequencies and various amplitudes and noted the corresponding dynamic supercooling obtained. The amount of supercooling was inversely proportional to the mechanical energy applied. Limitation of his experiment was the problem of the confinement of the liquid in the crucible without splashing and minimizing other unwanted modes of vibration. Tiller et al.8,9 did similar work on tin and Sn-Pb alloys using an electromagnetic stirring device. Their conclusions were that the magnitude of the magnetic field strength did not affect the amount of undercooling at which nucleation was initiated. While conclusive experimental results have been lacking to explain this effect of mechanical vibration on inducing nucleation, a number of theories have been proposed. Two of these theories are discussed below. 1) The Change in Melting:- Point Locally Due to the Change in Pressure (Clapeyron Equation). According to Vonnegut10 the most plausible explanation for the nucleation of a supercooled melt by cavitation is the effect of changing the melting point by a change in pressure. For materials where the volume decreases on solidification, an increase in pressure raises the melting point; for materials which expand on solidification, the melting point is raised for a decrease in pressure, i.e., rarefaction. Using the Clapeyron equation, the melting point of a metal can be calculated as a function of pressure. If it is assumed that the equation can also be used to calculate the temperature of nucleation of a supercooled melt as a function of pressure (i.e., the temperature of heterogeneous nucleation will increase with pressure at the same rate as the melting point), the amount of supercooling required for nucleation will be constant at all pressures as shown in Fig. 1. It is obvious that an isothermal change which results in an increase in melting point results in an equal increase in supercooling. This increase in supercooling may now be sufficient for nucleation. A pressure of 80,000 atm was calculated, using the Clapeyron equation, as the pressure required to increase the temperature of nucleation of nickel by 200°C. According to Lord Rayleigh,11 this very large pressure could be generated for a very brief period of time by the collapse of a cavity. This pressure wave is radiated in all directions from the collapsed cavity. If the temperature of the melt is slightly below its equilibrium melting temperature at atmospheric pressure, stable growth can follow; that is, once nucleation occurs, growth becomes the main driving force of the solidification process. This proposal has been extended to water which expands on freezing by assuming that nucleation occurs during rarefaction following the pressure pulse. This negative pressure pulse should follow immediately after the positive pressure pulse with its magnitude approaching the critical tensile strength of the liquid. The negative pressure developed during this period would raise the melting point of water and thus promote nucleation. Hunt and jackson12 have suggested this for water. Similarly, it could be postulated that bismuth which also expands on freezing could be nucleated during the negative pressure pulse. 2) Nucleation by a High-pressure Phase. An extension of the Clapeyron equation to systems where density decreased on freezing at atmosphere pressure has been proposed by Hickling.13 The phase diagram for water initially shows the well-known decrease in

Jan 1, 1969

By D. M. Milne, R. F. Stoner, W. Bakiewicz

The Lower Indus Basin deposits form a substantial unconfined alluvial aquifer of excellent quality. This aquifer contains basically highly saline groundwater, overlain in places by considerable amounts of fresh water whose origin can be traced to the present and historical courses of the River Indus. In the past, the main source of groundwater recharge has been direct river seepage and water tables were generally deep. Since the establishment of a perennial irrigation regime however, deep percolation from the increased irrigation supplies has predominated and the water table has risen to within a few feet of the surface. The river system itself is being progressively devel-oped for irrigation and very shortly a situation will arise where the entire low winter flow will be utilized. Fresh groundwater will then be exploited to supplement the surface supplies, and at the same time a drainage system, mainly based on wells will be used to stabilize the water table at safe depths. This paper discusses some of the methods used to ensure optimum economic development of the fresh groundwater and describes the general design philosophy evolved by the authors' firms for both irrigation and drainage wells, with respect to both well configuration and operation. Water quality standards are discussed insofar as they affect subsequent use of the groundwater. A few particular aspects are considered, including the values of the relevant aquifer constants, the construction of wells, and the determination of screen losses. A major problem solved during the Lower Indus Project investigations has been the design of wells to produce fresh water from layers overlying saline water. This solution is also discussed. The recent interest in the groundwater of the Lower Indus Plains of West Pakistan has two main aspects. Firstly, soil waterlogging and the associated salinity hazards caused by rising water tables are a serious problem in agriculture, and the ease with which drainage requirements can be met by aquifer pumping makes this an obvious solution. The second interest stems from the rapid intensification of irrigation agriculture and the growing shortage of water supplies. In the areas where the groundwater is of suitable quality it can be used immediately for irrigation at relatively small cost and with large financial returns. In addition, the aquifer forms a potential storage reservoir which could be recharged from high river flows and depleted by pumping at times of scarcity. SOURCE OF GROUNDWATER The Lower Indus plain is a relatively long and narrow alluvial basin on a north-south axis, some 23,000 sq miles in area, bounded by higher land to the east and west. One hypothesis to explain the distribution and quality of groundwater in the region1 presupposes a river system considerably larger than, but very similar to the present river. This is characterized by a single mainstream, with subsidiary overspill channels in times of flood, running along the crest of a self-deposited ridge and with an extensive advancing delta. Sea water is contained in the alluvial pores of the advancing delta, forming the native groundwater. This has been partially displaced from two sources, namely direct seepage of fresh water from the river channel in meander flood plain areas and percolation from flood spillage in cover flood plain areas. Percolation has been subject to much re-evaporation and is thus slightly saline even as it reaches the groundwater. The whole complex has been further affected by two main influences: evapotranspiration from the land surface that tends to salinize the groundwater and down-valley flow that tends to bring fresher water

Jan 1, 1970

By F. C. Bond

SIZE of grinding media is one of the principal factors affecting efficiency and capacity of tumbling-type grinding mills. It is best determined for any particular installation by lengthy plant tests with carefully kept records. However, a method of calculating the proper sizes, based on correct theoretical principles and tested by experience, can be very helpful, both for new installations and for guiding existing operations. As a general principle, the proper size of the make-up grinding balls added to an operating mill is the size that will just break the largest feed particles. If the balls are too large the number of breaking contacts will be reduced and grinding capacity will suffer. Moreover, the amount of extreme fines produced by each contact will be increased, and size distribution of the ground product may be adversely affected. If the balls added are too small, grinding efficiency is decreased by wasted contacts that are too weak to break the particles nipped; these largest particles are gradually worn down in the mill by the progressive breakage of corners and edges. Ball rationing is the regular addition of make-up balls of more than one size. The largest balls added are aimed at the largest and hardest particles. However, the contacts are governed entirely by chance, and the probability of inefficient contacts of large balls with small particles, and of small balls with large particles, is as great as the desired contact of large balls with large particles. Ball rationing should be considered an adjunct or secondary modification of the principle of selecting the make-up ball size to break the largest particle present. Empirical Equation In 19521,2 the author presented the following emerical equation for the make-up ball size: B - ball, rod, or pebble diameter in inches. F = size in microns 80 pct of new feed passes. Wi - work index at the feed size F. Cs - percentage of mill critical speed. S — specific gravity of material being ground. D == mill diameter in feet inside liners. K - 200 for balls, 300 for rods, 100 for silica pebbles. Eq. 1 was derived by selecting the factors that apparently should influence make-up ball size selection and by considering plant experience with each factor. Even though Eq. 1 is completely empirical, it has been generally successful in selecting the proper size of make-up balls for specific operations. But an equation based on theoretical considerations should be used with more confidence and have wider application. The theoretical influence of each of the governing factors listed under Eq. 1 was accordingly considered in detail, as described below, and a theoretical equation for make-up ball sizes was derived. Derivation of Theoretical Equation Ball Size and Feed Size: The basis of this analysis is that the largest ball in a mill should be just sufficient to break the largest feed particle into several pieces, excluding occasional pieces of tramp oversize. In this article the size F which 80 pct passes is considered the criterion of the effective maximum particle feed size. The smallest dimension of the largest particles present controls their breaking strength. This dimension is approximately equal to F. As a starting point for the analysis it is assumed that a 1-in. steel ball will effectively grind material with 80 pct passing 1 mm, or with F- 1000µ or about 16 mesh. The breaking force exerted by a ball varies with its weight, or as the cube of its diameter R. The force in pounds per square inch required to break a particle varies as its cross-sectional area, or as its diameter squared. It follows that when a 1-in. ball breaks a 1-mm particle, a 2-in. ball will break a 4-mm particle, and a 3-in. ball a 9-mm particle. This is in accordance with practical experience, as well as being theoretically correct. Confirmation of this reasoning is supplied by the Third Theory of Comminution," which states that the work necessary to break a particle of diameter F varies as F. Since work equals force times distance, and the distance of deformation before breaka4e varies as F it follolvs that the breaking force should vary as F½ These relationships are expressed in Table I, with a 1-in. ball representing one unit of force and breaking a 1-mm particle. This establishes theoretically the general rule used in Eq. 1 that the ball size should vary as the square root of the particle size to be broken. Ball Size and Work Index: The work input W required per ton" varies as the work index Wi, and the

Jan 1, 1959

By Zay Jeffries

IT is well known that many properties of a given metal vary with the size of grain or cell. For most industrial purposes, where high ultimate strength and high elastic limit are desired, the manufacturer tries to produce a. fine-grained structure. For some purposes, for example in transformer iron, a coarse-grained structure is desirable. The terms fine-grained and coarse-grained are used only in a relative sense. For example, a fine-grained cast copper might have grains 100 times larger than coarse-grained high-speed steel. Similarly, a fine-grained steel rail might have grains or cells 100 times larger than coarse-grained wire, which was made of steel from the same heat. In some metals, the change in grain size is more appreciable than the changes in any of the properties determined by the tensile test. Furthermore, in metals which have been annealed or subjected to high temperatures, grain size may prove to be a better indication of the life of the metal in use than the tensile test. One of the authors was able to supplant the tensile test with grain-size determinations in the control of. metal for a specific use, the latter determination being more indicative of. the life of the metal part than the tensile test. A grain-size determination with a tensile test is, of course, more valuable than either singly. At present there is a decided dearth of data on actual grain-size determinations in metals, and consequently a corresponding lack of ability to interpret the determinations after they are made. It would be desirable to have correlation made between grain size and properties of metals, comparing thousands of determinations. In this way, grain-size determinations might be interpreted in much the same manner that tensile tests are now.

Jan 12, 1915

By Michael Tenenbaum

RECENT researches on the reduction of iron ores have stimulated interest in the effect of increased pressures within the iron blast furnace. From a physicochemical viewpoint, it seems logical to suppose that the increased pressure would increase the rate of reduction of the iron ore within the furnace. This increased reaction velocity would in turn be reflected in an increase in the output of the blast furnace. Any sub-stantial effect of elevated blast-furnace pressures should accordingly be of both technical and economic interest to the operator. It is doubtful whether it is possible to increase the pressure in the stack of the blast furnace beyond two or three atmospheres without incurring severe mechanical difficulties. The benefits to be procured by elevations in pressure must therefore be restricted to pressures below two, or at the most, three atmospheres. Within this range it would seem both economically and mechanically feasible to modify the present design of the upper portion of the blast furnace to accommodate these more stringent operating conditions. This investigation was undertaken to study the effect of moderate pressures on the rate of reduction of iron ore with hydrogen. The use of carbon monoxide is contemplated for later work. Although any results that are obtained in the laboratory cannot be interpreted directly in terms of large-scale commercial operations, it is possible, through such an investigation, to ascertain the order of magnitude of the effects of higher pressure. A more quantitative conception of the effects that would result from the use of increased pressures in the modern blast furnace is needed.

Jan 1, 1939

By H. A. Kortmann, K. Koch, B. M. Grover, O. P. Burghardt

Addition of lime, especially when treating a hematite feed, sometimes results in swelling and a severe loss in strength during reduction in the temperature range 900° to 1100°C and in grain disintegration in the temperature range 400° to 600°C in a weakly reducing atmosphere. Swelling and/or grain disintegration of pellets is often the reason for trouble in blast furnace operation, resulting in a decrease in the rate of smelting. Following a brief description of the test procedures employed (measurements of swelling, pressure-reduction-softening and grain disintegration), some results are given which were arrived at when testing pellets from industrial plants. Results of systematic tests are presented, where the effect of the quantity and quality of the gangue upon the behavior during reduction of pellets is illustrated. Varying amounts of quartz and limestone were added to a high-grade hematite ore and, then, pellets produced in a pot grate. The relationships of the behavior of the fired pellets during reduction are presented, among others, in the ternary system CaO-FeOn-SiOP. Finally, an attempt is made to explain the reasons for swelling, loss in strength, and grain disintegration.

Jan 1, 1974

By Richard H. Jahns

Terminology and Basic Specifications Minerals and closely allied natural substances that are used for personal adornment, as raw stock for the fashioning of ornamental objects, or for other decorative purposes can be referred to collectively as gem materials. Industrial gem materials are those employed directly in various areas of industry. Substances that are manufactured for gem or industrial uses are termed synthetic gem materials; most of these correspond in composition and crystal structure to certain minerals found in nature. Imitation materials are those whose use is based upon their resemblance to substances of different composition and greater value. Gem stone is a designation commonly restricted to those materials suitable for personal adornment, and the complementary term ornamental stone is applied to the remaining materials not used industrially. This distinction is arbitrary and somewhat inconstant, owing to numerous geographic differences and historical shifts in end uses. Further, there is a complete gradation between the category of ornamental stones and that of dimension stone. Although many building and monumental stones are undeniably ornamental in their uses, treatment of such material is beyond the scope of this chapter. Substances that lose their identity in use are excluded from the gem category. Thus a transparent crystal of beryl would not be classed as a gem stone if it were treated as a source of beryllium metal. The adjectives precious and semiprecious have been employed as a means of distinguishing different gem materials on the basis of their recognized value. Diamond, emerald, ruby, and sapphire consistently have been regarded as precious stones, and in cut form they commonly are referred to as the noble gems. Opal, pearl, turquois, and chrysoberyl also have been classed as precious at various times. In recent years, however, it has been increasingly recognized that no satisfactory distinction can be based upon assessed value or any group of intrinsic properties, and at the present time all gem stones normally are referred to as precious. Indeed, the tendency is to extend this term to include many of the so-called ornamental stones, and the designation semiprecious has been essentially discontinued. Unlike most mineral commodities of economic value, gem materials are little altered in processing for end uses. Some stones are used in essentially their natural form, and most are reshaped and polished without significant changes in their fundamental properties. Thus the preparation of gem and ornamental stones is intended chiefly to emphasize or enhance desirable characteristics that are present initially. The worth of most synthetic gem stones is founded upon the identity between their major properties and the natural characteristics of the corresponding minerals, just as the pleasing qualities of imitation stones depend upon how well certain of their properties simulate those of more valuable substances. The most important specifications of gem materials represent four basic factors: beauty, durability, rarity, and portability. Beauty, whether expressed as splendor, "purity," attractiveness, or in some other way, is mainly a function of personal response to optical properties such as color, transparency, indices of refraction, and dispersion. Freedom from visible imperfections is important for transparent stones. For some substances beauty depends also upon asterism, opalescence, textural and

Jan 1, 1960

By A. H. Bromly

The small agricultural village of Santa Barbara, in the State of Guanajuato, has been the center of spasmodic tin-mining operations during recent years. The deposits, so far as I know, are unique, and a few brief notes may be of interest. Santa Barbara, about 12 miles west of San Felipe station on the Mexican National Railway, and at an elevation of 2,400 m. above sea-level, is situated in a flat valley extending east and west, and bounded north and south by broken rugged hills rising about 330 m. above the general level of the valley. Tin-mining operations have been conducted at the Queensland mine, near this village. A small concentrating-mill was erected and attempts were made to smelt the ore in reverbera-tory furnaces. The country-rock of the district for many miles is rhyolite of great thickness. There are some local variations due to differences in cooling, or subsequent alteration or impregnation; but throughout the whole district the principal constituents of the rocks remain the same, viz., free silica, orthoclase-feldspar and sanidine. The latter is especially marked. The crevices and joints of the rock are frequently filled with calcedonic and opalescent quartz. The ore-formation at the Queensland mine is of an unusual nature, consisting of cracks in the rhyolite which have been filled with clay. This phenomenon is largely developed over a width of from 80 to 100 m., in which the principal fissuring follows an approximate direction of north and south. The average dip ranges between 70 and 80 degrees. Some of the fissures hold sufficiently true for a time to afford the impression that they are due to dynamic causes, or earth-movement. This may be true of some of the north by south fissures; but considering

Jan 1, 1906

By Harold Rose

THE peak of production from the Appalachian natural gas field was apparently reached about 10 years ago, and the annual production from Pennsylvania, West Virginia and Ohio has now dropped to about two-thirds of the amounts obtained in 1916 and 1917. This .falling off in production has occurred in spite of vigorous efforts and large expenditures by producing companies to maintain or increase their output. The greatly increased costs of development and handling are reflected in the present average value of natural gas at the point of consumption, which value in the three states referred to, has more than doubled in the past 10 years. While it is true that the natural gas production of the United States has increased during recent years, this increase has come from southwestern states, particularly from Oklahoma, California, Louisiana and Texas, which states are now leaders in the production of this valuable but fleeting resource. Pennsylvania, originally the leading state in natural gas production, occupies sixth place in the latest (1924) U. S. Bureau of Mines statistics. West Virginia, which held first place for 15 years, has dropped to third rank, and Ohio is the seventh state in the order of natural gas production. Domestic and industrial demands for gas are constantly increasing everywhere in the United States. These growing demands have accentuated the difficulties in the localities with depleted natural gas supplies, and serious shortages have already occurred during winter months in the Pittsburgh district. Some relief has been obtained by placing restrictions on the industrial use of natural gas, and by increasing the price (which tends to reduce consumption). Yet these are but temporary expedients for checking gas consumption, which fail to solve the problem of adequate future supplies of gas.

Jan 10, 1926

By George Wolff

To THE average person, the purport of the items and figures on the balance sheet of a mining company are hazy and the real financial condition of the company is cloaked in obscurity. It is also likely that the satisfaction of the stockholder in his investment, as well as the determination of the prospective investor, rests more on the reputation of the company, his confidence in the management, the stated earnings per share and the dividends paid, than on a comprehension of the information contained in the balance sheet. The balance sheet of any company at best can only closely approximate the true financial condition as of a given date. Any item of asset, except cash, can probably be altered if another's viewpoint is applied. It is difficult and expensive to drill and explore an orebody to determine the entire tonnage of ore therein. Those inherent difficulties render it practically impossible to prepare an absolutely correct balance sheet of a mining company. It is possible, however, to present the facts of a mining company more intelligently and comprehensively than is usually done. With this, in mind, a balance sheet of an imaginary mining company is submitted, with supporting schedules. The same fundamental principles hold true whether we consider a large or a small company. Less complexity is involved in dealing with the affairs of a smaller company. Therefore, for this purpose, the assumed company operates but two mines and a smelting plant. It has been deemed advisable to introduce the features of the operation of a mine acquired prior to March 1, 1913 and one after that date. The presentation therefore is concerned in one mine with the appraisement as of March 1, 1913 and in the other with the appraisement by reason of discovery. A consequential factor pertains to dividends paid from depletion reserves and therefore free of surtax to the stockholders. While supporting schedules, when necessary, give details of certain items on the balance sheet, it seems desirable to give further explanations with respect to some of them. This seems to be important because some mining companies do not treat the items as indicated on the statements submitted.

Jan 1, 1928

By Verne E. Hooker, Robert M. Becker

The evaluation of in-situ rock stresses from measurements of the change in diameter of a borehole that is stress-relieved by overcoring has become a common practice. Generally the stress evaluation has been based on isotropic relations whereas rock in varying degrees is aniso tropic. A plane elasticity solution fora circular hole along one axis of an orthotropic medium has been adapted to evaluation of in situ stress. The necessary elastic parameters for this evaluation may conveniently be obtained from static tests on the overcore in conjunction with a consistent orthotropic thick-wall cylinder solution. The method and equations for the laboratory and field evaluations are summarized and illustrated. The determination of in-situ rock stresses from boreholc deformation measurements has gained widespread acceptance. Hastl measured the change in diameter of a borehole that was stress-relieved by drilling a larger concentric core; this overcoring method has produced a major portion of the field data for recent stress evaluations in rock. These evaluations are based on the elastic properties of the rock, which are conveniently determined by testing the overcore.6-8 The stress evaluation and the overcore testing procedures generally assume isotropic relations whereas rock is anisotropic in varying degrees. Recently, an anisotropic plane stress solution was adapted for evaluation of in-situ stress9 and a related plane strain solution was developed for the testing of stress-relief core. 10 The present paper unifies the stress evaluation and overcore testing procedures for a general class of anisotropic rock. Equations are presented for calculating the state of stress in a plane normal to the borehole, and a consistent thick-wall cylinder solution is given for determining the necessary elastic properties. In addition, a complete example of stress evaluation from laboratory and field data is presented, and the order of error in computing the state of stress from isotropic relations is illustrated. Several limitations of the results should be emphasized. All results assume small linearly elastic deformations of a homogeneous medium. Rock characterized by orthotropic, transversely isotropic, and isotropic elasticity is considered; however, the anisotropic results assume a special relation among certain elastic constants in addition to geometric orientation with respect to a plane of elastic symmetry. For some stress determinations in rock, these may be severe limitations. On the other hand, the anisotropic evaluation procedure involves little additional effort and is much more flexible than the corresponding isotropic evaluation procedure. The collection of field data is identical. The laboratory determination of elastic parameters is no more difficult and time-consuming for the anisotropic than for the isotropic model. Only the calculations require some additional effort for the anisotropic evaluation, and this additional effort is eliminated by use of a small computer. The flexibility of the anisotropic procedure is greater because more than two independent elastic constants are available in fitting rock data. For the special relationship assumed among orthotropic constants, five independent elastic constants are available for the determination of in-situ secondary principal stresses as compared with two independent elastic constants for the isotropic model. Much of the material presented is either directly from the literature or adapted from the literature through superposition or the usual transformations. Consequently, only enough detail for clarity of presentation is included. RELATIONS FROM ANISOTROPIC ELASTICITY Materials that have three mutually orthogonal planes of elastic symmetry are called orthotropic. By directing the axes of the coordinates perpendicular to these planes, Hooke's law for an orthotropic solid 11 is

Jan 1, 1970

By J. L. Gillson

Historically, rock deposits and sand deposits of titanium minerals came into production about the same time, although there may be some argument as to what is meant by production. Beach deposits of heavy minerals in India (Figs. 1-4) and Brazil (Figs. 5) were worked for monazite about the turn of century, but as there was then no market for titanium minerals, these were thrown away. The rock rutile deposits at Roseland, Va., Fig. 6, were worked to supply rutile for titanium chemicals and for coloring ceramics long before there was a titanium pigment business. The pigment industry started about the middle twenties, both in Europe and the U. S., and almost simultaneously the rock deposits at Ponte Vedra Beach near Jacksonville, Fla., were worked for titanium content. Since those days, production from both types of deposits has continued to grow at a rapid rate; many deposits of both types have been found, and reserves have grown to very large figures. In total tonnage of reserves, there is no doubt that the rock deposits are far ahead of the sand deposits; nevertheless there is a very large tonnage of commercial sands available. It is the quality of titanium mineral in the sand and the relatively lower costs of operating sand deposits that have kept them abreast, at least in annual tonnage produced, with the rock deposits. The principal titanium mineral used is ilmenite, but as soon as that mineral began to be sought as a titanium ore, it was obvious that there are ilmenites and ilmenites. Textbook ilmenite should have the composition FeOTiO2 and should analyze 52.6 pct TiO2 and 36.8 pct iron as Fe. The Indian ilmenite, for almost a generation the standard ore for manufacturing pigment in the U. S., was found to analyze about 60 pct TiO, and only 24 pct. Fe, and most of the iron is in the ferric condition. The whole process of pigment manufacture in the U. S. was built up on the use of a raw material of that grade, and the American chemical engineers who operate the pigment plants shuddered at the thought of using a rock ilmenite with 45 pct or so of TiO, and nearly 40 pct Fe. Intensive search was made around the world to find other deposits of rich black sand, like the Indian beaches, but although a few were found, there was some objectionable feature about each. A deposit in Senegal, south of Dakar (Fig. 7), was worked for a while, but an organic coating on the grains made attack by acid difficult. Modern practice would have included a scrubbing operation, in a caustic soda bath, to eliminate the organic coating. Brazilian deposits were numerous, but individually small, and shipping from them difficult. Deposits on the east coast of Ceylon had many attractive features, but the ilmenite analyzed only 54 pct TiO2 and could have been used only with a penalty. Sand deposits with 2 pct ilmenite, like those now worked in Florida, would not have been considered commercial ore, even if they had been known at that time. Most rock ilmenites are associated or mixed with hematite or magnetite, which accounts for the lower titanium and higher iron values than in the sand ilmenites. The Norwegians, English, and Germans, with cheap Norwegian rock ore at hand, learned to install in their pigment plants adequate capacity on the black side, as it is calltd, and counterbalanced the extra cost of plant, and larger amount of acid used, by the lower cost of ore. When World War II arrived, two of the largest pigment manufacturers in the U. S. had to learn how to use the Adirondack ilmenite, but one of them very gladly went back to sand ores when the Florida deposits came into large-scale production after the war. The other continues to use Adirondack ilmenite and finds it commercially attractive to do so. Rutile is not a raw material for titanium pigment manufacture by the sulfate process, since it is insoluble in sulfuric acid. In addition to its small consumption in chemicals and ceramics it began to be used in quantity in welding rod coatings. With the outbreak of World War 11, and the tremendous need for welding rods in shipbuilding and other structural steel construction, rutile suddenly became in heavy demand. The sand deposits on the eastern shore of Australia (Fig. 8A) which had been worked in a small way since 1934 were brought into production, and some stream placers in Brazil were worked and rutile concentrates shipped to American

Jan 1, 1960

By W. Ivanick, Frank W. Glaser

ORDERING reactions of Fe-Si compositions in the a region have been mentioned in a number of review articles.7-8 However, little is known in regard to the critical temperature, T at which this transformation takes place. Greiner and Stoughton,' in a study of the constitution of iron-rich Fe-Si alloys below 800°C, placed the onset of order at a composition of about 4 to 5 wt pct (7.5 to 9.5 atomic pct) Si. They observed that at this composition the ductility and strength begin to decrease sharply and brittleness develops. During X-ray studies they observed only the appearance of superlattice lines for compositions containing 12.5 atomic pct Si. More recent X-ray diffraction studies- etected the appearance of superstructure lines (corresponding to the calculated positions of [Ill] and [ZOO] reflections) for alloys containing 9.5 atomic pct. This fact confirms Greiner and Stoughton's original hypothesis in regard to the onset of order at compositions containing about 7.5 to 9.5 atomic pct Si. The physical properties.of Fe-Si alloys containing more than 10 atomic pct Si have been studied only to a limited extent. Their extreme brittleness and difficulties of sample preparation render them of little industrial use in comparison with alloys in the 0 to 10 atomic pct Si range. The properties of Fe-Si alloys reported to date are mostly properties measured at room temperature on ordered compositions It appears that because of experimental difficulties fully disordered alloys could not be produced in the critical composition ranges and their properties could therefore not be reported. In Figs. la, lb, and lc, a summary of physical properties of Fe-Si alloys, measured during recent investigations of the influence of silicon additions to iron, is presented. No mention of determination of the critical temperature, T for disorder could be found in the literature. The following is an account of work done in an attempt to: 1) produce the disordered state of various Fe-Si alloys and establish the critical temperature, T for disordering, and 2) compare the electrical and thermal conductivity characteristics of the ordered and disordered phases as functions of composition. Materials, Preparation of Test Samples, and Testing Methods Fe-Si alloys with high silicon contents (about 10 atomic pct) are usually prepared using powder metallurgy techniques. This method insures accurate compositions and high purity through the use of protective atmospheres during solid state diffusion alloying of the mixed powders."" Iron and silicon powders were mixed in the desired proportions, cold or hot pressed, and then homogenized at 1150°C for periods ranging from 6 to 12 hr under a protective getter and in a hydrogen atmosphere. Fully homogenized samples were quenched in water from various and successively higher temperatures to establish the critical temperature for disorder. Temperatures were determined using Pt—Pt-Rh thermocouples in contact with the test samples. Test bars were approximately 0.6x1.25x7.5 cm. Electrical measurements were made at room temperature by determining the potential drop over a 1.5 cm span for a current of 10 amp. Densities of the samples were determined by immersion. High temperature electrical resistivities were measured using previously described techniques.'" Specific heat conductivity up to 200°C was measured by heating one end of the test bar and water-cooling the other; heat conductivity was then obtained by subtracting the rate at which heat entered the water from the

Jan 1, 1957

By D. A. Rice, M. C. Fuerstenau

Flotation data indicate that sulfonate and amine adsorb physically on manganese oxide; oleate also adsorbs physically if the zero point of charge is sufficiently high. Chemisorption of oleate occurs on pyrolusite between pH 8 and 9, and possible mechan- isms of collector adsorption involving metal ion hydrolysis are discussed. The flotation characteristics of most metal oxides1-15 have been examined with standard collectors such as fatty acids and sulfonates. From these studies it would appear as though electrical phenomena are important when shorte r-chained collectors are involved. For example, Iwasaki, Cooke, and Choi9 have shown that dodecyl amine and dodecyl sulfate float iron oxide above and below the zero point of charge, respectively. In another system, D.W. Fuerstenau and Modi10 have shown that dodecyl sulfate, dodecyl

Jan 1, 1969

By Charles Catlett

My attention having been called several years ago to the possibility of increasing the yield of coke per ton of coal, as obtained in certain bee-hive ovens, I called the attention of the Institute to the matter at the New York meeting in 1899. I have had occasion since to witness the indifference with which this matter is ordinarily considered, and, on the other hand, the marked improvement which comes from even a most perfunctory effort in that direction; and I am almost tempted to say that nothing in connection with the coking of coal is at present so important. There are other factors, which are fundamental and necessary; but they have been considered already ; while the question of yield is, as a rule, absolutely ignored, or some figure is taken as the result of a special burning of one or two ovens, and the assumption is made that the whole battery is yielding at that rate. The uncertainty of this assumption I may illustrate by saying that, in my own observation, the burning of a single oven in a battery frequently gives, by actual weight, as much as 69 or 70 per cent. of merchantable coke, while the battery as a whole was yielding only an average of 63 per cent. Only from careful records, checked at least once a month, of the weight of the coal going to the ovens and the weight of coke produced, can any correct idea of the actual yield be gained, or methods be modified so as to produce good results; yet, as a rule, the crudest estimates are usually the only available data. In many cases there are reasons for this loose practice. The coal is mined by measurement; the payment for drawing the ovens is in proportion to the amount of cod which goes to them; and, not uncommonly, the royalty is so much a ton of coke, not of coal. The mine-manager is tempted to increase the volume of coal mined which is paid for as a ton, and to do the same thing in charging the ovens. Nor does he

Jan 1, 1903

By George W. Maynard

At the Pittsburg meeting of the Institute, May, 1879, I made the first announcement in America of the results obtained by Sidney Gilchrist Thomas and Percy C. Gilchrist, in their efforts to eliminate phosphorus in the manufacture of steel in the Bessemer converter and the open-hearth furnace.' The first published statement in an American newspaper was a reprint of the first Thomas-Gilchrist paper,which I sent to J. C. Bayles, at that time the editor. At the suggestion of many friends who were present at the Pittsburg meeting, and others who are cognizant of the beginnings of the process in the United States, and for the information of the younger generation of the iron and steel fraternity, I have thought it eminently proper that the history of the process in detail should be given where I made the first announcement and where the process has had its largest development. During my residence in England, from April, 1873, to February, 1879, and as a member of the Iron and Steel Institute since 1874, I was present at many of its meetings during those years. At nearly every meeting the burden of discussion was the question of the use of pig-iron containing phosphorus in the manufacture of steel by the Bessemer or the open-hearth process. In my biographical notice of Thomas: read at the New york meeting of the Institute, February, 1885, I show. that the working-out of the process was not a hap-hazard or accidental inspiration, but the culmination of many years of investigation of experiments and theories which had been carried on aid advanced by chemists and metallurgists at home and abroad.

Jan 1, 1911

ROBERT WOOLSTON HUNT, who joined the Institute in 1874, served twice as its President, was made Honorary member in 1919, in whose honor the Hunt prize and medal were established, and who, within a month, had received the Washington Award, died at his home in Chicago on July 11, 1923. Robert Woolston Hunt was born Dec. 9, 1838, in Fallsington, Buck County, Pa., His father, Dr. Robert A. Hunt, of Trenton, N. J., graduated from Princeton College in the class of 1824, and later from the Uni-versity of Pennsylvania. His mother was Martha Lancaster Woolston. Mr. Hunt spent several years learning the practical side of iron making in the roll-ing mills of John Burnish & Co. at Pottsville, Pa. Later he took a course in analytical chemistry in the laboratory of Booth, Gar-rett & Blair, upon the completion of which he entered the employ of the Cambria Iron Co. at Johns-town, and for them, on Aug. 1, 1860, established the first laboratory in Amer-ica forming an integral part of an iron or steel organization. In the fall of 1861 he en-tered the United States military service, and was in command of Camp Curtin at Harrisburg, serving as mustering officer for the State of Pennsylvania, with the rank of Captain. In 1864 he assisted in recruiting Lambert's Independent Mounted Company, and was mustered into active ser-vice as a sergeant, having tossed up with a friend who had also participated in recruiting the company, as to which one should receive a Lieutenant's commission. Upon being mustered out of service, Mr. Hunt returned to the employ of the Cambria Iron Co. which sent him to the works at Wyandotte, Mich., where experiments with the new Bessemer process of making steel were being conducted. He was placed in charge of those works in July, 1865, and so continued until May, 1866, when the Cambria company called him back to Johnstown to take charge of their steel business, intending at once to begin the erection of a Bessemer steel plant. The plant, however, was not built for several years, and in the meantime the Cambria company began the rolling of steel rails for the Pennsylvania Railroad, using ingots of Bessemer steel produced by the Pennsylvania Steel Co. at their Steelton works. Mr. Hunt had charge of the steel for these operations; and in August, 1867, the first steel rails made in America on a commercial order were rolled.

Jan 8, 1923

By George Rice

SOME form of longwall mining is generally used in Continental Europe; also in Great Britain where the coal is weak and friable, or the coal bed provides material for pack walls and filling, or where the bottom is soft and squeezes up easily, or the roof is pliable, or the bed is thin and brushing provides building material, or the thick multiple seams are mined in layers, such as the 24-ft. seam at Weymss, Scotland or the 10-yd. pitching bed near Coventry, England. Pillar methods have been retained in British fields where the coal beds are from 5 to 9 ft. (1.5 to 2.7 m.) thick, and free from thick partings or binders or without draw slate, which would provide waste rock for pack walls, or where the roof is hard and requires shooting to bring down and also the bottom or floor is comparatively hard. The room-and-pillar system, by which probably 95 per cent., of the coal of the United States is produced, is now found in only a few mines in Wales, where it is known as the pillar. and .(single) stall method. The American room-and-pillar system is not equivalent to the bord-and-pillar or stoop-and-room system. The essential difference between the bord-and-pillar and the average American room-and-pillar system is probably due to natural conditions. The majority of the coal mines in the United States have a depth of less than 700 ft. (213.m.). The bulk of the coal mined in Europe comes from a depth of more than 1200, ft., while the deepest mines in Great Britain, Belgium, and France reach 3500 to 4000 ft. In deep workings wide spans are impracticable, even with the use of abundant timber; accordingly, as the mines in the United States become deeper the rooms will become narrower and the pillars thicker, and the bulk of the coal will be extracted on the retreat. It is generally conceded, without reference to the cost of production, that the larger the pillars left on the first mining, the more thoroughly can the, coal be extracted. In Europe complete extraction is generally compelled either by the lessors or by the governments; whereas in this

Jan 2, 1921