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By Wilbur Nelson

Little drilling has been done in Kentucky and Tennessee in 1923, with the exception of the Kettle Creek area on the Kentucky-Tennessee line. Oil prices fluctuated between $2.86 and $1.36 Per bbl. for light oil. The Kentucky production was slightly over 8,000,000 bbl., being 1,000,000 bbl. less than during 1922. Tennessee production was approximately 8500 bbl., being a decrease of 1000 bbl, over. 1922. During 1923, the first producing well was brought in, in Clay County, Tenn. Production in Tennessee will undoubtedly increase considerably in 1924, and Kentucky production will probably increase slightly. THE year 1923 has been a dull period in oil development in Kentucky and Tennessee, with the exception of the Kettle Creek area along the Tennessee-Kentucky line, in Clinton and Cumberland Counties, Kentucky, and Clay County, Tennessee. The price of oil during the year has been so low that wildcatters have shown little interest in drilling. At the beginning of the year, light oil sold for $2.31 per barrel and medium oil for $2.06 per, barrel. Prices increased to the highest point for the year in February, when light oil brought $2.86 per barrel and medium oil, $2.61. In April, prices had declined 35 c. while in May an additional decline of 25 c. went into effect. During June, prices declined 20 c. per barrel; during July, 45 c. per barrel; remained stationary during August and September, and dropped during November and December to the low price of $1.36 per barrel for light and $1.21 for medium oil. At present, an increase of 29 c. has brought the price of light oil up to $1.55, and indications are that the price will continue to increase during the coming months. It can readily be seen that during 1923 the price of oil was such as to prevent any drilling except that necessary to hold leases and to maintain production enough to take care of overhead expenses.

Jan 3, 1924

By C. W. Schultz, R. B. Tippin

Statistical screening is so named because a particulate feed is separated into "size" intervals on the basis of the probabalistic behavior of the individual particles (i.e., their probability of passing the screen) rather than by physical restraint. To illustrate this statement, consider a conventional screening operation. A particle which is larger than the openings in the screen cloth is physically restrained from passing through and is repeatedly reflected from the surface until it is discharged as oversize. Of those particles smaller than the screen openings, the very smallest tend to pass the screen near the feed end while those nearer in size to the screen opening tend to pass at a greater distance from the feed end. Alternatively, we could say that the particles with high probability of passing through the screen are concentrated near the feed end while particles of low probability of passing are dispersed from the feed end. This tendency to segregate is the basis for size separation by statistical screening. The purpose of this note is to: 1) Show that the distribution of material passing an array of screens can be described by a modified binomial distribution. 2) Demonstrate that an efficient size separation can be made by this technique. 3) Relate some of the more important screening parameters to the efficiency of the size separation. The Distribution Function The distribution of undersize products along the length of a screen can be derived with the help of a few simple definitions and assumptions. 1) First, we must assume that each particle has some probability of passing a screen on a single presentation. Further, we must assume that this probability is constant and is not affected by the position of the particle on the screen. It must also be acknowledged that several factors enter into this probability. Among them are the "action" of the screen, the slope of the deck, the shape of the openings, etc. The relationship between these factors, however. is not the subject of this paper. 2) To simplify calculations, all particles falling within a size range on the Tyler scale, e.g., —x + x/ mm, are assumed to have the same probability. 3) It is assumed that particles which do not pass the screen on a presentation are reflected a unit distance forward, i.e., the direction of material travel. A unit distance is defined as the distance between the centers of consecutive wires in the screen cloth.

Jan 1, 1971

By H. D. Colson, Robert A. Rapp

When a binary alloy is oxidized to form simultaneously an internal oxidation zone and an external scale which grows at a constant rate, the oxidation should reach a condition of steady shale, in which the internal oxidation zone is maintained at a time-independent thickness. Assuming an ideal model for the steady-state oxidation, the diffusion equation is solved for the steudy-state concentration profiles for oxygen and alloying- element in the metal. A theoretical prediction of' the steady-state thickness of the internal oxidation zone is made in terms of the experimental parameters. FOR the oxidation of binary alloys, Rhines, Johnson, and Anderson,1 as well as Mask: have derived equations describing the kinetics for the simultaneous formation of an internal oxidation zone and an external scale which was assumed to mow according to a parabolic rate law. Experimental determinations of internal oxidation zone thicknesses beneath a Cu2O scale for Cu-Be and Cu-A1 alloys3,4 have shown some deviation from Maak's theoretical expressions, because the ideal model assumed for the calculation is not met experimentally. Some pure metals, such as columbium and tantalum oxidize at elevated temperatures and high oxygen pressures to form nonprotec-tive external scales which thicken at a constant rate. If such a base metal is alloyed with a solute element which may form a more stable oxide than the solvent oxide, then internal oxidation of solute may occur beneath the growing external scale. A review of the internal oxidation of alloys has been published recently.5 The purpose of this paper is to analyze theoretically for a binary alloy the kinetics for the simultaneous formation of an internal oxidation zone and an external scale which grows at a constant rate. Consider an alloy A-B in which A is the more-noble solvent and B is the less-noble solute. The solute oxide is assumed to precipitate in internal oxidation as stoichiometric particles of BO, which are assumed to have a negligible solubility product in the A matrix. It is further assumed that the diffusion of oxygen through the zone of internal oxidation is not impeded by previously precipitated BO, particles. The specimen is assumed to be sufficiently thick so that the solute content in the center of the specimen does not change. For consistency: the following notation of Maak2 has been used:

Jan 1, 1967

By Bruce Chalmers, D. C. Larson

Aluminum crystals with longitudinal-axis orientations of (111) . (110), and (100) were deforined in tension and annealed. The conditions of deformation were controlled so that the re crystallization nuclei originated in either the heavily deformed regions at saw cuts {artificial nucleation) or in the lightly deformed matrix (spontaneous nucleation). The artificial-nucleatioln experiments showed that in lightly deformed (110) and (100) crystals low-angle twist boundaries are most mobile, while in (111> crystals and heavily deformed (110) and (100) crystals high-angle tilt boundaries with near (111) rotations are favored. The spontaneous-nucleation experiments showed the existence of preferred orientations in the (111) crystals. The nonrandomness of the grain orientations is quantitatively determined through a comparison with the results which would he obtained from a randowl set of grain ovientations. PREVIOUS recrystallization studies have been performed on single crystals deformed in tension.1 7 The crystals used in these studies usually had random tensile-axis orientations and the extent of deformation was not a primary consideration. The present study concerns the recrystallization of single crystals with tensile-axis orientations of (Ill), (110), and (100). The emphasis of this work is on the influence of the tensile-axis orientation and the degree of deformation on both the nucleation and growth processes. The multiple-slip orientations were chosen because secondary slip or slip intersection promotes nucleation.1,5,8 These crystals recrystallize at lower strains than the crystals which are oriented for single slip. Also, the greatest variation in deformation behavior is exhibited by the multiple-slip orientations. The stress-strain curves for crystals with tensile-axis orientations of (111) are higher than the stress-strain curves for poly-crystals, and the stress-strain curves for crystals with tensile-axis orientations of (100) are lower (at large strains) than the stress-strain curves for the crystals which deform initially in single slip.g The recrystallization nuclei originated in either 1) the homogeneously* deformed matrix of the crys- tals or 2) the heavily and inhomogeneously deformed regions at saw cuts. The nuclei will be referred to hereafter as spontaneous and artificial nuclei, respectively. The two terms do not imply a difference in the nature of the nuclei; they imply simply a difference in the mode of introduction of the nuclei. During spontaneous nucleation very few (always less than ten) grains nucleate, while during artificial nucleation large numbers of grains nucleate. Only a fraction of the artificially nucleated grains penetrate very far into the deformed matrix during annealing. The grains that penetrate the farthest into the deformed matrix will be referred to as the dominant grains. EXPERIMENTAL PROCEDURE The thirty-five crystals used in this investigation were grown from the melt in milled graphite boats at a rate of 1.6 cm per hr. The crystals had dimensions of approximately 6 by 12 by 80 or 6 by 6 by 80 mm and the aluminum was of 99.992 pet purity. The as-grown crystals were annealed at 610°C for 24 hr and furnace-cooled. They were then heavily etched and electropolished in a solution of five parts methanol to one part perchloric acid. The crystal orientations were obtained by back-reflection Laue photographs and were accurate to ±2 deg. The tensile-axis orientations were (loo), (110), and (111). Two of the side faces of the (111) crystals were (110) lanes. The (110) crystals had both {100) and {110) side faces and the (100) crystals had (100) side faces. The crystals were deformed at a strain rate of 0.003 per min. Shear stress and shear strain were obtained by multiplying and dividing the tensile stress and strain, respectively, by the Schmid factor, m. For the (111) crystals m = 0.272 and for the (110) and the (100) crystals m = 0.408. The Schmid factor is effectively constant during deformation for all orientations. The deformed crystals were sawed into 1-in.-long specimens while the crystals were totally enclosed in a graphite boat. The sawing was performed very carefully in order to limit the plastic deformation to the sawed regions. The specimens were electropolished in the solution mentioned above to remove the sawed-end deformation as well as controlled amounts of surface material. A special stainless-steel grip was used to hold the specimens during the electropolishing treatment. The gripping faces were flat, with no teeth, to prevent the introduction of extraneous de-

Jan 1, 1964

By H. S. Chiang, D. F. Lu, S. S. Peng

INTRODUCTION The most important element in a successful lingual mining is a good roof control. The modern longwall mining employs hydraulic powered supports for roof control at the face area. The application of hydrau¬lic powered support requires the knowledge of over¬burden strata behavior for proper selection of sup¬port type and capacity. Failure to do so could lead so serious loss. There are several methods available for determining the required support capacity (1-3). While these methods are simple for application, they do not include the complicated roof behavior observed in longwall mining. As research progresses and operational experience accumulates (4,5), the concept about the designing and selection of powered support improves. The design of a longwall powered support consists of three major phases: 1. structural integrity and stability of the powered support, 2. external loadings induced by the movements of the overburden strata, and 3. interaction between the support, roof and floor. Phase 1 involves structural analysis (5) and full-sized testing (6) of the supports. Its validity is limited by the accuracy of the assumed external loading because of the uncertainty about the actual loading underground. The third phase includes the reaction of the support and the floor to the movements of the overburden strata and vice versa. Among the three phases, the second phase concer¬ning the external loading seems to be the least known because of the complicated behavior of the roof strata. There are many unresolved problems. For example, does the main roof break periodically and cause periodic roof weighting in the face area? If so, are there any rules governing its behavior? How does the roof load on the support canopy! Finally, how can one determine the required support capacity and select a proper type of support to meet a certain roof behavior? In order to answer those questions, underground instrumentation and observations were performed at 4 longwall panels in 3 separate mines for the past two years. This paper summarizes the current findings. PANEL LAYOUTS AND EQUIPMENT EMPLOYED The three mines selected are all located in West Virginia; two in northern and one in southern West Virginia. As shown in Table 1, seam conditions (i.e. seam, depth and thickness) and panel layouts are different among the three mines. The most significant difference in equipment is the face powered supports. Three mines used three different types of shield; 2-leg caliper, 2-leg lemniscate, and 4-leg lemniscate chock-shield. (Fig. 1) UNDERGROUND INSTRUMENTATION AND OBSERVATION PROGRAM Two events were instrumented in each observed longwall face: one was the hydraulic pressure (resistance) of the powered supports and the other was the canopy load distributions. In addition, the gob caving conditions were visually observed and recorded. Leg and Support Resistances One or two automatic Weksler Pressure Recorders were installed at the designated shield support,. In most cases, the daily charts were used to record the pressure variations in both the front or the rear legs (for the 4-leg shield), or in both the leg and the fore-pole ram (for the 2-leg shield). The recorded pressure w a s then converted to load or resistance by multiplying it by the cross-sectional area of the hydraulic leg or canopy ram piston. Fig. 2 shows the typical pressure-recorded charts for the 4-leg and 2-leg shields in a 23-24 hour period. The support resistance is the summation of the resistance in each of all the legs for that support. Generally, the resistance of the fore-pole ram will not be considered in determining the capacity of the support because of its rather small vertical compo¬nent force at the tip of the fore-pole. Canopy Load Distribution External load distribution on the canopy as exer¬ted by the roof was monitored. The measurements employed 12-14 pieces of pressure cells (6-inch square) that were uniformly arranged in two rows on the canopy. After support setting, the pressure changes in the cells were monitored at various stages of the mining (supporting) cycle while the support leg pressures were recorded continuously by the pressure recorders. Based on the calibration chara¬cteristics of each pressure cell as performed in the laboratory before and after each underground test, the cell pressures were converted to actual loadings. From these load measurements the canopy load distri¬butions and the relations between measured canopy loadings and support leg resistances were determined. Accordingly, the supporting efficiency of the shield support can be determined.

Jan 1, 1982

By J. Chipman, M. N. Dastur

The importance of dissolved oxygen as a principal reagent in the refining of liquid steel and the necessity for its removal in the finishing of many grades have stimulated numerous studies of its chemical behavior in the steel bath. From the thermodynaniic viewpoint the essential data are those which determine the free energy of oxygen in solution as a function of temperature and composition of the molten metal. A number of experimental studies have been reported in recent years from which the free energy of oxygen in iron-oxygen melts can be obtained with a fair degree of accuracy for temperatures not too far from the melting point. Certain discrepancies remain, however, which imply considerable uncertainty at higher temperatures; also several sources of error were recognized in the earlier studies. It has been the object of the experimental work reported in this paper to reexamine these sources of uncertainty and to redetermine the equilibrium condition in the reaction of hydrogen with oxygen dissolved in liquid iron. The reaction and its equilibrium constant are: H2 (g) + Q = H2O (g); K1 _ PH2O / [1] Ph2 X % O Here the underlined symbol Q designates oxygen dissolved in liquid iron. The activity of this dissolved oxygen is known to be directly proportional to its concentrationl,2 and is taken as equal to its weight percent. The closely related reaction of dissolved oxygen with carbon monoxide has also been investigated:3,4,5 co (g) +O = CO?(g); K _ Pco2___ [2] K2= pco X % O [2] The two reactions are related through the wat,er-gas equilibriuni: H2 (g) + CO2 (g) = CO (g) + H2O (g); K2 = PCO X PH2O [3] PH2 X PCO2 and with the aid of the accurately known equilibrium constant of this reaction, it has been shown5 that the experimental data on reactions [1] and 121 are in fairly good, though not exact, agreement. Experimental Method Great care was taken to avoid the principal sources of error of previous studies, namely, gaseous thermal diffusion and temperature measurement. The apparatus was designed to provide controlled preheating of the inlet gases and to permit the addition of an inert gas (argon) in controlled amounts, two measures found to be essential for elimination of thermal diffusion. A known mixture of water vapor and hydrogen was obtained by saturating purified hydrogen with water vapor at controlled temperature. This mixture, with the addition of purified argon, was passed over the surface of a small melt (approximately 70 g) of electrolytic iron in a closed induction furnace. After sufficient time at constant temperature for attainment of equilibrium the melt was cooled and analyzed for oxygen. GAS SYSTEM A schematic diagram of the apparatus is shown in Fig 1. Commercial hydrogen is led through the safety trap T and the flowmeter F. The catalytic chamber C, held at 450°C, was used to convert any oxygen into water-vapor. A by-pass B with stopcocks was provided so that the hydrogen could be introduced directly from the tank to the furnace when desired. From the catalytic chamber the gas passed through a water bath W, kept at the desired temperature by an auxiliary heating unit, so that the gas was burdened with approximately the proper amount of water vapor before it was introdvced into the saturator S. All connections beyond the catalytic chamber were of all-glass construction. Those connections beyond the water bath were heated to above 80°C to prevent the condensation of water vapor. After the saturator, purified argon was led into the steam-hydrogen line at J, and finally the ternary mixture was introduced into the furnace. THE SATURATOR The saturator unit comprised three glass chambers, as shown in Fig 1, the first two chambers packed with glass beads and partially filed with water and the third empty. Each tower had a glass tube with a stopper attached for the purpose of adjusting the amount of water in it. The unit was immersed in a large oil bath, which was automatically controlled with the help of a thermostat relay to constant temperature, ± 0.05ºC, using thermometers which had been calibrated against a standard platinum resistance thermometer. The performance of the saturator over the range of experimental conditions was checked by weighing the water absorbed from a measured volume of hydrogen; the observed ratio was always within 0.5 pct of theoretical.

Jan 1, 1950

By Charles Clayton

CAST-STEEL runners, while not interesting from a commercial standpoint, furnish valuable material for microscopic study. Foley1 found not only the usual ingot structure, but zones of Widmannstattian structure, which he explains as due to differential crystallization. The writer, in examining nickel-steel runners, found a type of pro-eutectoid ferrite that is unusual and most probably new. The runner in question was 8.75 cm. (3.5 in.) in diameter and analyzed at its center 2.69 per cent, nickel and 0.350 per cent. carbon and at its edge 2.62 per cent. nickel and 0.359 per cent. carbon. The macrostructure, which is the usual type, is shown in Figs. 1 and 2, which are a cross-section and the longitudinal surface, respectively. Slight segregation of ferrite can be seen near the center of the segment and at other points near the edges of polyhedral grains. Stead's reagent brings out the dendritic structure within the grains as shown in Figs. 3 and 4. This texture is what would be expected of such a steel. Figs. 5 to 10 show the peculiar swirls of pro-eutectoid ferrite, the finding of which prompted this paper. These swirls, or eddies, if such a term can be applied, occur in all parts of- the runner and in all parts of the individual grains. I he tipper left-hand figure shows this ferrite at a grain boundary; the middle left-hand figure shows the inner portion of a grain; and the lower left-hand figure shows these swirls at the edge of a grain which is also the outer edge of the runner. Straight-lined ferrite associated with swirls is found in the areas of Figs. 11 and 12. The texture of the carbon-bearing constituent of the runner is shown to- some extent in Fig. 12, the pearlite being very fine or sorbitic. Higher magnification does not bring out any new features either in the peculiar ferrite or in the pearlite. Figs. 13 and 14, at 240 diameters, show the curved and straight-lined ferrite.

Jan 3, 1920

By Samuel S. Wyer

The art of natural-gas compressing is now over 25 years old, and has grown at practically the same rate as the increase in domestic natural-gas consumers. There are now over 200 natural-gas compressing stations in North America, aggregating more than 320,000 hp. of compressor capacity and representing a property value of more than $22,000,000, and compressing more than 85 per cent. of all the gas used. The age and magnitude of the art make it evident that the use of gas compressors is a recognized integral part and universal custom of the natural-gas business. The public also is not without its rights and vital interests in this problem. Approximately 1,700,000 domestic natural-gas consumers in North America are dependent upon gas compressors for their natural-gas service. That is, if the use of compressors were to be prohibited, the majority of these consumers would be unable to secure adequate natural-gas service. Each consumer represents between four and five persons, and it therefore follows that the comfort and well-being—as far as natural-gas service is concerned—of at least 8,000,000 persons would be affected if the use of natural-gas compressors should be prohibited. These 1,700,000 consumers have invested, in services, house piping, fixtures, and appliances, an average of about $90 each, or an aggregate of $153,000,000, which is much more than the companies' investment in gas compressors. Furthermore, in all cases where the rate paid by the consumers is fair to the gas company—considering the value of the service rendered by the gas company—the consumers are entitled to continued service and protection of the investment they have made on the faith that the gas service would be continued in the future. Since "Customs adopted and acquiesced in, if not in conflict with federal or State legislation, have the force of positive law,"l and ('Courts will take notice of whatever ought to be generally known withxn the limits of their jurisdiction,"2 there ought to be no question as to the un-

Jan 1, 1916

By B. M. Bird

ONE of the problems studied by the U. S. Bureau of Mines in cooperation with the University of Washington has been the re-treatment of table middlings. Hydraulic classification has given the best results. Instead of using the classifier to prepare the feed to the tables, as has been the practise for many years in ore dressing, it is used to separate part of a middling product that cannot be re-treated successfully on the table. Middlings are desirable in a table plant because the separation according to specific gravity is incomplete; that is, in the products of the table there is always an overlapping of free particles of coal and refuse that can be distinguished with heavy liquids such as are used in the float-and-sink test. The causes of this overlapping may be divided into two groups, those outside the table separation itself and those inherent in the table separation. The first of these groups includes, among other causes, the following: (1) poor adjustments of the table, (2) overloading or underloading the table, (3) irregularities in the quantity and quality of the feed, (4) irregularities in the power and water supply, and (5) a vertical danc-ing motion of the table deck due to weak foundations. These causes of overlapping are largely unnecessary in a properly designed and constructed plant; but whenever they are present they materially affect the character of the middling product, and hence the problem of re-treatment. The other group includes three additional causes: (6) small differences in specific gravity between coal and impurities, (7) a size-specific-gravity relationship of the coal and refuse particles in the table feed such that the sizing action of the table opposes the separation according to the specific gravity, and (8) shapes of particles that prevent a separation. The sixth, small differences in specific gravity, may be a very important reason for an incomplete separation. The change in specific gravity from coal to shale is always gradual and the designation of one part as coal and another part as refuse is arbitrary. If some specific gravity, such as 1.45, is chosen as the dividing point between washed coal and refuse,

Jan 1, 1928

By Franz R. Dykstra, R. F. Tatnall

A ferroalloy is defined as: "An alloy of iron with some element other than carbon used as a vehicle for introducing such an element into the manufacture of steel. The element may alloy with the steel by solution or, as the carbide, -neutralize the harmful impurities, by combining with them and separating from the steel as flux or slag before solidification."a The most commonly used ferroalloys are: ferromanganese, ferrochrome, ferrosilicon, ferrovanadium, ferromolybdenum, ferronickel, ferrotungsten, ferrocolumbium, and ferrophosphorus. Concern here is with the sources and markets of the major metallic alloying elements. The discussion is further restricted to those - ferroalloy ores, the principal use of which is in the production of steel. Phosphorus, silicon, and tungsten, for example, are ferroalloys of some economic significance, but other uses dwarf their alloying applications. Thus, this discussion will deal with the ores of managanese, chromium, nickel, molybdenum, vanadium, and columbium. As shown in Table 14.1B.1 ferroalloys serve to- facilitate processing and impart to the finished steel such characteristics as tensile strength, ductility, low-temperature impact, hardenability, high-temperature strength, corrosion resistance, toughness, and abrasion resistance. In the steel-making process, ferroalloys may be added in the charge, in the molten bath near the end of the finishing period, in the ladle, or in the molds. Timing of the addition depends on the effect of the addition on the temperature of the molten metal, the formation and elimination of reaction products, and the protection of subsequent additions from oxidation. MANGANESE General Approximately 95% of the manganese ore sold is consumed by the steel industry, either as ferromanganese or as manganese metal. Manganese is the universal alloy, without which good steel cannot be made. It serves as a scavenger, removing phosphorus, sulfur and, to a degree, oxygen from the melt. As an alloying material it makes steel hard, tough, and resistant to wear. Approximately 6 kg of manganese is required for every ton of common steel produced by United States steelmaking practice. The remaining 5 % of the manganese ore consumed consists of special or so-called "chemical" ores, including high-grade pyrolusite ores, electrochemically reactive "battery ores," and a number of special varieties, as well as those otherwise regarded as "metallurgical ores."

Jan 1, 1976

By Milo W. Krejci

The plant and quarries of the Inland Portland Cement Co. are located at Metaline Palls, Wash., about 128 miles north of Spokane, on the Pend Oreille river, and within 10 miles of the Canadian border. The plant is one of a chain of cement plants controlled by the Lehigh Portland Cement Co., of Allentown, Pa. Work on this plant was commenced in June, 1910, and was completed in April, 1911, but on account of some delays in the construction of the power plant, the works were not put into operation until August, 1911. Since then it has been in continuous operation and has furnished the Inland Empire with a high grade of cement. Fig. 1 is a view and Fig. 2 is a plan of the works. Water Power Development.—The source of the water supply for the hydro-electric plant of the Inland Portland Cement Co. is Sullivan lake, a body of water approximately 5.5 miles long and' 0.5 mile wide, situated in the Kaniksu Forest Reserve. A wooden crib dam was installed so as to raise the elevation of the lake 25 ft., and the additional storage capacity was taken up by the diversion of Sullivan creek into the lake by means of a wooden flume, 0.25 mile long, 6 by 8 ft. in cross-section, and a ditch 1,200 ft. long. The water from the lake follows the natural water course to an intermediate storage reservoir, which was created by the construction of an earthen dam. This pond was used in connection with the saw mill which was installed at this point. Besides the earth dam, there was also built a crib, at the point shown in the sketch map, Fig. 3. This crib dam is 35 ft. high. From this point the excess water follows the natural course of Sullivan creek to the Pend Oreille river. The water is taken from this intermediage storage by means of a 6 by 8 ft. wooden flume approximately 13,000 ft. long. This flume is built up on the side hill, piling being driven most of the way on

Jan 1, 1914

THERE is probably no phase of basic open-hearth steelmaking that is more of an art and less of a science than mold and pouring practice. It varies widely from plant to plant; consequently it is obviously impossible to set down in detail in this chapter the many ramifications of practice that are now used. The discussion that follows therefore attempts only to give an outline of the more common methods, with examples sufficient to indicate some of the variations in procedure, and to emphasize a few of the many underlying problems that face the open-hearth operator at the pouring platform. MOLD DESIGN Ingot molds are hollow castings of considerable wall thickness. Despite this, non-uniformity in wall thickness should be strictly avoided to prevent the setting up of unnecessary stresses. Molds are cast in various forms, sizes, and shapes depending on the type of steel, and on whether the ingot is to be withdrawn from the mold or the mold from the ingot. The selection of ingot-mold material, which is usually cupola or blast-furnace iron containing approximately 3.5 per cent carbon, 1 per cent silicon, 0.9 per cent manganese, 0.2 per cent phosphorus, and 0.07 per cent sulphur, is governed by numerous considerations that affect mold life. (1) Factors in addition to wall thickness that should be given primary consideration at the foundry are casting skin, shrinkage stresses, and casting temperature. Factors in Mold Design. In use, mold thickness is of greatest importance, for it must be heavy enough to freeze the molten metal rising in the mold and to reduce the temperature of the steel below that of the melting point of cast iron to prevent the inner surface of the mold from melting, which would result in the welding of the ingot to the mold. The skin of an ingot, which solidities immediately upon contact with the relatively cold walls of the mold, must soon become thick enough to contract on the fluid mass and pull away slightly from the sides of the mold. The separation of the ingot from the mold depends entirely on the heat taken up by the mold from the steel.

Jan 1, 1944

By J. D. Forrester

SINCE the introduction of the first percussion rock drill in about 1850, drilling equipment has undergone a progressive evolution so that a wide range of special classes of machines now are marketed by several manufacturers. Essentially all of the drilling machines presently employed for blasthole drilling are of the hammer-drill type. They are classified as hand-held hammers (jackhammers, sinkers, etc.), drifters (light or heavy), and stopers and are of many sizes; varying in weight and rotative and blowing power. Most makes of drill machines fundamentally are the same in principle of operation as they are designed to strike a blow on the end of a drill rod in the manner of early-day hand drilling. However, beyond this common characteristic, the similarity among different makes ceases. In choosing between similar machines of one make or another, it is important to remember that rock drill manufacturers have not attempted to standardize drill parts, the result being that a mine operator must approach, because of replacement parts, standardization by limiting the makes of machines used. The present-day selection of any particular drilling machine is dependent fundamentally on its inherent ability to make a hole, or series of holes, as speedily and as cheaply as is possible. Two categories of controlling elements become apparent in any analysis devoted to making the choice of a drill, namely: 1-Conditions under which the equipment is to be used, such as mining method practiced, type of labor available, character of rock material. 2-Physical characteristics or features of the given drilling machine. Regardless of the controls of class 1, above, the elements of this category are fixed by several factors. All of them are interrelated and, though some often are more important in one case than in another, an unfavorable variation of any-others being essentially equal-is sufficient to discredit the optimum applicability of the particular drill machine. The relative factors in the selection of a machine are as follows: First cost-This is not generally an important item on a comparative basis as the marketing of drill machines is highly competitive and most makes are priced closely.

Jan 1, 1952

By Charles T. Holland

Miners have observed for many years that as pillars are removed in mining operations the pillar adjacent to those mined out frequently shows evidence of being compressed. Although this has been known for a long time, little or no information is available concerning how far from the mining area the pillars are affected in this manner, or how great is the magnitude of the compression, or concerning the manner or characteristic way in which the compression acts, as the mined-out area approaches the pillar under consideration. The writer made a study of pillar action in the Mona mine of the Arkwright Coal Co., partly to get some quantitative data on the deformation of coal-mine pillars. The study was not completed but it seems best to present the information that was obtained. Method of Measurement Fig. 1 shows the details of the apparatus used for measuring the deformation. This is permanently attached to the pillar. Fig. 2 shows the details of the gauge that records the deformation. This gauge is removable and was used to measure the deformation at all the stations. The recording dial is graduated in thousandths of an inch and records measurements up to one inch. Fig. 3 shows how the removable gauge is placed in the instrument. When installing the instrument in the side of the pillar, all loose coal near by was removed and a trench 4 to 6 in. deep, and approximately 1 ft. wide, was dug in the pillar. In this trench near the top and the floor, the holes holding the instrument were bored. These holes were then partly filled with thin cement and the rods were inserted. The remaining space in the holes was tamped full of cement. At this time, by carefully manipulating the permanent rods and clamps and adjustable pin, it was possible to adjust the instrument so that the removable gauge would stand true and measure any vertical deformation between the top and bottom rods. Fig. 4 is a photograph of one of the instruments installed, showing the removable gauge in place.

Jan 1, 1938

By Charles T. Holland

Miners have observed for many years that as pillars are removed in mining operations the pillar adjacent to those mined out frequently shows evidence of being compressed. Although this has been known for a long time, little or no information is available concerning how far from the mining area the pillars are affected in this manner, or how great is the magnitude of the compression, or concerning the manner or characteristic way in which the compression acts, as the mined-out area approaches the pillar under consideration. The writer made a study of pillar action in the Mona mine of the Arkwright Coal Co., partly to get some quantitative data on the deformation of coal-mine pillars. The study was not completed but it seems best to present the information that was obtained. Method of Measurement Fig. 1 shows the details of the apparatus used for measuring the deformation. This is permanently attached to the pillar. Fig. 2 shows the details of the gauge that records the deformation. This gauge is removable and was used to measure the deformation at all the stations. The recording dial is graduated in thousandths of an inch and records measurements up to one inch. Fig. 3 shows how the removable gauge is placed in the instrument. When installing the instrument in the side of the pillar, all loose coal near by was removed and a trench 4 to 6 in. deep, and approximately 1 ft. wide, was dug in the pillar. In this trench near the top and the floor, the holes holding the instrument were bored. These holes were then partly filled with thin cement and the rods were inserted. The remaining space in the holes was tamped full of cement. At this time, by carefully manipulating the permanent rods and clamps and adjustable pin, it was possible to adjust the instrument so that the removable gauge would stand true and measure any vertical deformation between the top and bottom rods. Fig. 4 is a photograph of one of the instruments installed, showing the removable gauge in place.

Jan 1, 1938

By Ira B. Joralemon

The Ajo copper district is in the heart of the Arizona desert, near the western boundary of Pima county. Gila Bend, the nearest railroad point, is 43 miles north of the camp, and the little Mexican border town of Sonoita is 30 miles south. Between Gila Bend and Sonoita, Ajo is the only settlement save for one or two small cattle ranches and the uncertain villages of nomadic Papago families. The Little Ajo mountains rise a few hundred feet above wide desert valleys, beyond which are lava mesas. The camp itself lies at an elevation of about 1,900 ft. above sea level, in a little basin on the east side of the range, separated from the open desert by low hills. In the center of the basin the brilliantly iron and copper stained rocks of Copper mountain rise 150 ft. above the village. Many varieties of cactus and low desert bushes cover the hills, while in the large valleys there are clusters of mesquite, paloverde, and ironwood trees which furnish limited amounts of fairly good fuel. A scanty supply of water is obtained from wells in the desert and from- prospect pits in the camp. The nearest abundant source of water is the Gila river, 50 miles north of Ajo. (See Fig. 1.) , History of the Ajo District The Ajo, next to Senta Rite, New Mexico, is claimed to have been the first copper district in the Southwest worked by Americans. In the '60's or even earlier high-grade native copper and cuprite ore was mined from shallow surface workings and hauled by bull team 400 miles across the desert to San Diego. From there it was carried in sailing ships to Swansea, Wales. Later, the ore was hauled to Yuma, floated down the Colorado river to the Gulf of California, and shipped to Swansea. Bat-infested old workings and the massive axles of the bull carts are shown to prove the truth of the stories. From this early period until the beginning of the present century, the Ajo district was worked only in a casual and intermittent way. With the increased interest in copper mining during the past decade, the brilliant

Jan 1, 1915

By W. M. Weigal

ECONOMIC deposits of barite occur in Missouri in two main districts. The most important, the Southeastern or Washington County district, is in the southeastern part of the state, mostly in Washington County with a small extension into the northeastern part of St. Francois County. Roughly, it embraces nine townships or about 200,000 acres and accounts for from 80 to 90 per cent. of the state production. The Central district, in the central part of the state, includes most of Morgan and Miller counties, and parts of Camden, Benton, Moniteau and Cole counties. The area embraced is much larger than the Washington County district, but the deposits are more infrequent. The location and general area covered by the two districts is shown in Fig. 1. There is another small area in southern Franklin County but so far its production has been of small consequence. It might be considered as a northward extension of the Washington County field. An area in the central part of Hickory County has also produced a small amount. The outline of the districts as shown indicates the general area in which deposits have been found. The total area covered by deposits of barite is only a small part of this, as, in the Central district especially, there may be miles of supposedly barren areas between ore-bearing areas of an acre or two in extent. The date when barite was first mined in Missouri cannot be definitely stated, but the industry probably dates back to about 1860. The occurrence of barite was known long before this, as it was encountered in the lead mines in Washington County and mentioned by several earlier writers.' Missouri is reported as producing 8000 tons in 1882 so it is evident the industry was well under way in the seventies in order to have arrived at this substantial production. By 1893 Missouri production had increased so that it ranked with that of Virginia, then the largest producer. It increased until in 1914 Missouri accounted for about 65 per cent. of the total United States output. Recent Production of Barite Table 1, compiled from Mineral Resources, published by the United States Geological Survey and Bureau of Mines, gives the production of

Jan 1, 1929

By John W. Roberts

The Federal income tax law permits as a deduction in determining net income "in the case of mines,... a reasonable allowance for depletion and for depreciation of improvements, according to the peculiar conditions in each case, based upon cost, including cost of development not otherwise deducted: Provided, that in the case of such properties acquired prior to Mar. 1, 1913, the fair market value of the property (or the taxpayer's interest therein) on that date shall be taken in lieu of cost up to that date." The law provides for a reasonable allowance but the regulations provide for a definite method of computation that possesses more the virtue of simplicity than that of reasonableness. Article 210 of Regulations 45 reads: "When the cost or value as of Mar. 1, 1913, or within 30 days after the date of discovery of the property shall have been determined, and the number of mineral units in the property as of the date of acquisition or valuation shall have been estimated, the division of the former amount by the latter figure will give the unit value for purposes of depletion, and the depletion allowance for the taxable year may be computed by multiplying such unit value by the number of units of mineral extracted during the year." This is simple, but is it reasonable? The intent of the law is clearly that the cost of the ore in the ground shall be considered as part of the cost of the same ore when removed and sold. The regulations fail to carry out that intent, because of the unwarranted assumption that all the ore in the mine cost the same amount per ton. This is what gives the Treasury Department's method its simplicity, but at the same time robs it of its reasonableness because the assumption is contrary to fact. It is self-evident that a ton of ore near the surface or exposed by present workings is worth much more than a ton buried so far down in the earth that it cannot be removed for many years. If no improvements in mining methods were expected, the value of deeply buried ore would be further reduced by the greater expense that would be required to

Jan 1, 1922

By Walter S. Weeks

THE subject of the flow of air through a mine under the influence of natural ventilation has proved an elusive one. To my knowledge only cut and try algebraic solutions have been used where multiple paths are involved. The object of this paper is to present a method for the graphical determination of the flow induced by the natural ventilation pressure (N.V.P.) alone or in combination with a fan. The N.V.P., measured in pounds per square foot, induced by two columns of air between two horizontal planes at different elevations, is equal to the difference in weight of the two columns, assuming each to have a cross-sectional area of 1 sq. ft. Pressure in pounds per square foot is converted into inches of water by dividing by 5.19. Methods of calculating the N.V.P. have been discussed in the literature. In this paper, for simplification of exposition, the change in pressure due to the weight of superimposed air will be assumed directly proportional to the length of the column. The actual methods of calculating the N.V.P. do not affect the solutions to be presented. It must be realized that the precision to be obtained from the small- scale, inked diagrams presented is not as great as would be obtained from large-scale, penciled curves that would be used in practice. The accuracy of the curves does not warrant reading "volumes" closer than the nearest thousand cubic feet per minute. CASE I. DUCTS IN SERIES The simplest type of problem is to determine the flow when two shafts are connected by a crosscut, as shown in Fig. 1. In this example, the resistance of the shafts is to be taken as 0.0001 in. of water for a flow of 1000 cu. ft. per minute for a length of 1000 ft. The resistance of a crosscut is to be taken as 0.0005 in. of water for a flow of 1000 cu. ft. per minute. The small pressures necessary to cause changes in velocity will not be considered, because the accuracy of other obtainable data does not warrant their consideration.

Jan 1, 1938

By A. H. Lee

Discussion of the paper of H. P. HOWLAND, printed in Bulletin No. 111, March, 1916, pp. 627 to 650. A. H. LEE, Buffalo, N. Y. (communication to the Secretary*). The statements and results recorded in Mr. Howland's paper can be accounted for by the use of coke with a low sulphur content (0.60 and under). This allows the furnaces to be run on a small volume of acid slag, which it is presumed was done in Mr. Howland's practice. H. A. Brassert has admitted that this was the largest factor in low fuel consumption in his practice. Where the coke contains from 1.00 to 1.25 per cent. of sulphur the situation is quite different and I doubt whether the results mentioned by Mr. Howland could be obtained under such circumstances. ` N. M. LANGDON, Mancelona, Mich. (communication to the Secretary + ).-Gruner's ideal working (which is but a statement of the law first enunciated by Bell) as given in Gordon's Translation of Studies of Blast Furnace Phenomena, "supposes that the reduction of the oxide of iron is effected by the CO only, without intervention of solid carbon." This ideal, if reached, would mean theoretically perfect or 100 per cent. efficiency of reduction. The workings of the large number of furnaces as given in Bell's and Gruner's great works do not show that this standard of perfection ever was reached. This problem has been worked out by the writer for a number of furnaces, and his results, as well as those given by others, all show a failure to reach perfection. Some furnaces, however, approach nearer to perfection than others. The "ideal" as given by Gruner was not stated as a solved problem in blast-furnace practice, but as an ideal to the attainment of which furnace managers should strive. A study of the factors involved from every angle is interesting and instructive and the paper on this subject contributed by H. P. Howland is no exception to the rule. This article is fairly open to criticism, not because of the unique method of presenting the subject, but for the reason that, as the writer conceives, erroneous assumptions are taken which lead naturally to incorrect conclusions.

Jan 7, 1916