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By H. M. Kraner, E. B. Snyder

This paper shows that volume stability, low porosity and decreased pyroplasticity are desirable for blast-furnace linings, partitularly for the hearth. It shows further that a hot load test is a valuable means of testing the fusion or softening behavior of a refractory at operating temperatures. The effect of carbon monoxide on commercia1 blast-furnace refractories in their as-received condition and after refiring is reported, showing that many commercial blast-furnace refractories disintegrate badly but that refiring decreases the effect and certain special refractories are now available which are almost free of the tendency. Factors Affecting Choice of ._ Refractories If one were to depend entirely upon experience or trial in the selection of clay refractories, the solution of a problem would be a slow process. Furnace campaigns are so long and attended by so many variables that it is difficult to draw fine distinctions within a reasonable period of time in regard to the quality of refractories used. It could be said, of course, that if service results do not yield the proper information for intelligent choice, there is no difference in the quality of brick being considered, but this is not necessarily true. Conventional ceramic tests are designed to give the ceramic engineer information as to the refractoriness and firing tempera- . ture, which he in turn internrets in terms of volume stability, ability to withstand load at high temperature (pyroplasticity), permeability, chemical stability and re-. sistance to the action of slag. In most refractory applications . several of these factors are involved. In a particular instance, one requirement may be predominant, while in others some other factors may be of primary importance. In this country, clay refractories are generally made from mixtures of plastic and flint fire clays. The process of firing a clay refractory is one of slowly melting the constituents, and during this operation its porosity is gradually reduced. The process is arrested, of course, in its incipient Stages and the Progress is determined by measuring slight changes in porosity, volume, bulk gravity, etc., that have taken place. Often the temperatures to which clay refractories are exposed in service are higher than those employed in their firing. However, if the manufacturer were to fire them at such high temperatures the brick would distort. Therefore it would be difficult for the manufacturer to fire all of his clay refractories to temperatures at which they are to be used, although in many cases it would be desirable from the user's point of view if he would do so. The porosity of clay refractories decreases by firing, to a minimum limit, after which a further increase in the temperature causes expansion in certain clay constituents and bloating in others. Both may occur in the same refractory containing two such clays.

Jan 1, 1944

By R. L. Huntington, D. T. MacRoberts, Charles R. Hetherington

Since the equation of continuity governing transient flow of gases through porous media cannot be integrated mathematically into a simple usable expression free from series terms, empirical and approximate equations have been advanced relating cumulative production and future production rates from gas reservoirs with time. In order to check the accuracy of these empirical equations and to obtain an insight into the initial conditions pertinent to the solution of the fundamental equation of continuity, experimental work has been conducted on a vertical 2-in. tube, 91.6 ft. high, packed with unconsolidated Wilcox sand. The results presented in graphical form show the accuracy of the approximate relations. Data from unsteady-flow runs on the 2-in. experimental tube are presented graphically, and are applied to obtain an equation for this specific case, and are offered as an aid in future investigations of this basic equation. The close agreement between equations calculated for steady-state flow and the experiments indicates that the decline of actual gas reservoirs may be treated by an application of these equations without serious error. Introduction The steady-state flow of gas through a porous material—that is, flow wherein the boundary conditions do not vary with time—is subject to analytical solution. Muskat1 has shown that the general differential equation is simply the La Place equation in P2. Direct solutions may be at once obtained. The unsteady state, on the other hand, is not analytically reducible. Since nearly all phenomena of practical interest in gas production involve flow in the unsteady rather than in the steady state, it has been necessary to derive approximate Solutions to the general equation. The validity of such approximations has not hitherto been adequately tested. In the experiments to be described a simple, closed, linear system is substituted for the real complex gas field. The general results, however, are independent of the geometry Of the system' The problem studied is the rate of production, and the accompanying variations in pressure, when one end of a linear porous system is opened to the atmosphere and permitted to produce to exhaustion. An approximate solution to the general differential equation, based upon the assumption of an "infinite series of steady states," is then applied to the results to determine the validity of this method of dealing with actual reservoirs. Theoretical Considerations The fundamental differential equation governing the flow of a gas through a Porous medium can be derived easily from the equation of continuity, Darcy's law, Boyle's law and a statement of the thermo-dynamic nature of the flow. For isothermal flow it is: d2P2/dx2 + d2p2/dy2 + d2p2/dz2 = 2fµ/K dP-dT

Jan 1, 1942

By R. L. Huntington, D. T. MacRoberts, Charles R. Hetherington

Since the equation of continuity governing transient flow of gases through porous media cannot be integrated mathematically into a simple usable expression free from series terms, empirical and approximate equations have been advanced relating cumulative production and future production rates from gas reservoirs with time. In order to check the accuracy of these empirical equations and to obtain an insight into the initial conditions pertinent to the solution of the fundamental equation of continuity, experimental work has been conducted on a vertical 2-in. tube, 91.6 ft. high, packed with unconsolidated Wilcox sand. The results presented in graphical form show the accuracy of the approximate relations. Data from unsteady-flow runs on the 2-in. experimental tube are presented graphically, and are applied to obtain an equation for this specific case, and are offered as an aid in future investigations of this basic equation. The close agreement between equations calculated for steady-state flow and the experiments indicates that the decline of actual gas reservoirs may be treated by an application of these equations without serious error. Introduction The steady-state flow of gas through a porous material—that is, flow wherein the boundary conditions do not vary with time—is subject to analytical solution. Muskat1 has shown that the general differential equation is simply the La Place equation in P2. Direct solutions may be at once obtained. The unsteady state, on the other hand, is not analytically reducible. Since nearly all phenomena of practical interest in gas production involve flow in the unsteady rather than in the steady state, it has been necessary to derive approximate Solutions to the general equation. The validity of such approximations has not hitherto been adequately tested. In the experiments to be described a simple, closed, linear system is substituted for the real complex gas field. The general results, however, are independent of the geometry Of the system' The problem studied is the rate of production, and the accompanying variations in pressure, when one end of a linear porous system is opened to the atmosphere and permitted to produce to exhaustion. An approximate solution to the general differential equation, based upon the assumption of an "infinite series of steady states," is then applied to the results to determine the validity of this method of dealing with actual reservoirs. Theoretical Considerations The fundamental differential equation governing the flow of a gas through a Porous medium can be derived easily from the equation of continuity, Darcy's law, Boyle's law and a statement of the thermo-dynamic nature of the flow. For isothermal flow it is: d2P2/dx2 + d2p2/dy2 + d2p2/dz2 = 2fµ/K dP-dT

Jan 1, 1942

By James K. Stanley

PREFERRED orientation, particularly in irons and low-carbon steel, is a phenomenon that is both of considerable importance and theoretical interest. At times it is a liability and at other times an asset. In deep-drawing operations, preferred orientation is extremely undesirable because it necessitates trimming operations and increases scrap loss. Furthermore, oriented metals are likely to crack more easily than randomly oriented material because the forming of symmetrical and unsymmetrical shapes out of preferentially oriented metals and alloys increases the probability of rupture in regions of high stress, owing to differential ductility.1 In the electrical industry, on the other hand, preferentially oriented silicon irons are in demand for transformer application because of the improved magnetic permeabilities in the rolling direction.2 This phenomenon of preferred orientation in soft magnetic materials has been responsible for the tremendous improvement in transformer design and construction in recent years. Directional properties in metals have received a great deal of attention in the metallurgical literature,3-9 in which much of the information reported about orientation has been obtained by conventional methods, mechanical properties and X-rays. A wealth of data has been obtained and often at a great expenditure of time. The recent introduction of the magnetic torque measurements on ferromagnetic materials provides a very simple and rapid method of evaluating orientation. The judicious use of the method can supply important in- formation, but it should be borne in mind that such studies must be supplemented by X-rays, optical methods,10 and etch and pressure figures" to get a rational under- standing of the orientation phenomena. No claims are made that the magnetic torque of evaluating orientation will reveal the behavior of the metal in deep drawing. For this purpose the accepted tests for deep-drawing quality must be used. For a good drawing 'stock, which must have low impurities, desirable grain size, ductility, strength, hardness, good surface, random orientation, and good recrystallization behavior, it cannot be expected that any simple test will be satisfactory for appraising all the factors. The only factor on which the magnetic method can be used is orientation, which is, of course, of considerable importance. When directionality can cause a 10 to 15 per cent change in tensile strength and a IOO to 300 per cent change in tear length,12 it goes without saying that a reliable as well as rapid method for detecting these changes would be desirable. There are certain inherent advantages in testing orientation by the torque method: The method is rapid with the proper type of equipment; no elaborate preparation of specimens is required-it is only necessary punch or machine a disk of the sheet material and insert it into the testing device.

Jan 1, 1943

By E. J. Duggan

THE mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. DESCRIPTION OF ORE The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chalcopyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the dose association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. HISTORY Construction of the first milling unit was started in 1917 and it was operated during parts of 1918 and 1919. In the postwar years, there was almost no demand for the product and production was at a standstill until 1924. The Company's research campaign to develop the use of the metal began to show results in that year. From 1926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and' 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for molybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill 10,000 tons per day, but it was not long before 12,500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. To increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. CRUSHING There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. 1) and No. 2, of 15,000 tons capacity (Fig. 2).

Jan 1, 1942

By J. E. Owen

EXPERIMENTS on twenty-six drilling muds were carried out in an effort to establish a correlation between the apparent viscosity values for these muds as obtained with the Marsh funnel and the Stormer viscometer. This required a calibration of both instruments in which the density of the calibrating liquids was taken into account. Corresponding to the Marsh-funnel viscosity values that were obtained at 1500 c.c. in, 1 qt. out, equivalent Stormer viscosity values were obtained at a Stormer cylinder test speed of approximately 1100 r.p.m. Further, a linear relation between the logarithms of the Stormer cylinder test speed and the resulting apparent viscosity was found to exist for the muds tested for 200 to 1000 r.p.m. The significance of the slope of this linear relationship is discussed as it applies to the field use of drilling muds. It was concluded that Marsh-funnel viscosity measurements are expressible in absolute units that may be correlated with Stormer viscosity measurements, provided these are made in the manner described in the paper. INTRODUCTION The Marsh funnel is used almost universally to determine the viscosity of drilling muds in the field. The Stormer viscometer is commonly used to determine the viscosity of drilling muds in the laboratory. With the Marsh funnel, the time required for one quart of mud to flow out of the funnel originally filled with 1500 c.c. is often taken as the measurement of viscosity. With the Stormer instrument, drilling mud is tested at a cylinder speed of 600 r.p.m. and by means of calibration data the apparent viscosity is expressed in absolute units. Neither of these instruments is an absolute viscometer, and apparent viscosities obtained with either are influenced by the arbitrarily standardized conditions under which the measurements are made. It is disconcerting to the man in the field to be told that laboratory measurements of the viscosities of a group of muds differ widely, when he has already observed that they have identical viscosities when tested in the Marsh funnel. Drilling muds are not perfect fluids, therefore it is not surprising to find a lack of correlation between the apparent viscosity of a mud as determined with the Marsh funnel and that determined with the Stormer viscometer. The work described here was undertaken to develop a correlation between Marsh-funnel and Stormer viscosity measurements. To a considerable degree this has been accomplished. This study has resulted not only in a better understanding of the two viscosity instruments and the data obtainable from them, but has given considerable information regarding the viscosity characteristics of drilling muds. EXPERIMENTAL PROCEDURE Calibration of Harsh Funnel and Stormer Viscometer The first step in the experimental procedure was the calibration of the Marsh funnel and the Stormer viscometer with

Jan 1, 1941

By F. H. Allen, R. P. Roe

The performance history of a volumetric gas condensate reservoir is presented. Curves depict the pressure-production relation and illustrate the phace behavior of the reservoir fluid. This performance is compared to calculated performance. This Bacon Lime Reservoir was discovered in 1941 and has been produced to field separators and gas sales. Reservoir performance has been observed by periodic bottom hole pressure sur. veys and companion sample analysis of well effluents, in addition to gas and condensate production data. , The pressure-production history demonstrates that the reservoir has produced under volumetric control. The "retrograde" behavior of the reservoir fluid is illustrated by decreasing butanes-plus content of the well effluent, followed by revaporization as the reservoir pressure approaches abandonment. Calculated performance predictions are compared to actual performance and some discussion offered on apparent discrepancies. INTRODUCTION Since the late 1930's, many gas condensate reservoirs have been developed, the retrograde theory of phase behavior has been generally accepted, and a number of cycling programs have been initiated. Numerous technical papers have since appeared in the literature dealing with the theory and application of cycling the gas condensate reservoirs. Some cycling operations have now advanced to a stage where it is possible to determine performance characteristics to a fair degree of accuracy. Field performance data for gas condensate reservoirs operated by pressure depletion have not been so extensively reported in the literature, although their probable behavior has been discussed in explanation of the retrograde theory. It is the purpose of this paper to compare calculated theoretical reservoir performance with field observations and well stream analyze; which have been obtained throughout the practical producing life of a small volumetric gas condensate reservoir. The reservoir under study is the Bacon Lime Zone of a field located in East Texas. The Bacon Lime occurs in the Lower Glen Rose Formation of Cretaceous Age. This Bacon Lime Reservoir is encountered at an average depth of 7600 ft (7203 ft subsea elevation) and comprises approximately 3100 acres located on a structural high with 80 ft of productive closure. The producing formation consists of approximately 50 ft of dense crystalline fossiliferous dolomite of generally low permeability with typical wide variation in both permeability and porosity, the more permeable stringers measuring 30 to 40 millidarcys. The average porosity is extimated to be 10 per cent with a connate water saturation of approximately 30 per cent of the pore space. The reservoir temperature was found to be 220°F and the original pressure was 3691 psia at -7200 ft elevation. The re ervoir was discovered in November, 1.941, and is now essentially depleted, Throughout most of its productive life, from the standpoint of cumulative withdrawals, the reservoir has been produced by five wells. During its latter stages of depletion, 16 wells were completed in the Bacon Lime.

Jan 1, 1950

By Guillermo V. Borquez

INTRODUCTION The basic factors that influence the selection of a mining method are the geometry of the ore deposit and the strength of the ore and surrounding rock. These factors usually determine the ore recovery and dilution of the deposit under analysis. For the purpose of this paper we will consider a generalized case of a block caving system and the sequence of the analysis needed to optimize ore recovery. The analysis presented is intended for the preliminary planning stage. The critical factors involved in the caving process are identified and the methods for evaluating the suitability of a particular deposit to the caving method of mining are outlined. The mine design is approached by identifying the two most important, dimensions: the size and spacing of the draw points. These dimensions are estimated from a simulation model which takes into account the ore recovery, the incremental value of the ore recovered, and the safety conditions of the operations. The model estimates the number of draw points required to extract efficiently and safely the broken ore from the caving area. This number is found in the range where the net incremental value of the ore is maximized and the probability of pillar failure does not exceed the limit which is acceptable for a safe mining operation. A mining method and its design are in a continuous process of optimization. This process is unique for each deposit because of the many parameters involved; therefore, a back analysis of the operations should be performed when possible. These analyses will indicate if the ore body responds in a predictable manner to the parameters contained in the geological, fragmentation, and draw point models. SELECTION OF ALTERNATIVE MINING METHODS The selection of a mining method best suited to define the minable ore reserves of a given deposit involves consideration of the geometry of the ore deposit and the strength of the ore and surrounding rock. These two factors are usually sufficient to determine the ore recovery and dilution of the deposit. After analyzing the geological reserves and defining the limits of the deposit, either an open pit or an underground mining method can be selected to define the minable ore reserves. The sequence of steps to follow in the selection of a mining method is presented in Figure I. From an economic point of view, the strip- ping ratio is the determining factor for the selection between an underground or open pit method. In turn, the stripping ratio is a function of the grade distribution and con- figuration of the ore body. In order to obtain a preliminary estimation of the stripping ratio, a parametric analysis could be applied, such as the approach used by Professor E. P. Pfleider (1972) and summarized in Table 1.

Jan 1, 1981

By R. J. Borg, C. E. Birchenall

Vapor pressures of eight compositions of Mg-Cd have been determined by the Knudsen effusion technique. Measurements are made at several temperatures for each alloy and the results interpolated to a common temperature for the calculation of the activities. It is not possible to obtain values for the partial molar enthalpies because of the narrow temperature range spanned by the measurements. The results show less precision than those obtained for pure Cd which are also reborted. These alloys are extremely susceptible to corrosion which provides a possible explanation for the lack of precision. Nevertheless, the values of the activity are in good agreement with the results of electromotive force measurements previously reported for the Mg-Cd system. MAGNESIUM and cadmium are mutually soluble in all proportions in the solid state and form a single phase which has the hexagonal close-packed structure. X-ray investigations of this system1,2 have revealed a rapid variation in the relative dimensions of the unit cell occurring at approximately 50 at. pct. The rate of change of the c/a ratio with composition increases markedly for alloys containing more than 50 at. pct Cd. The purpose of this investigation was to determine the activity of Cd in the neighborhood of 50 at. pct and to discover whether or not the thermo-dynamic properties undergo a similar rapid change. The activities are calculated from vapor pressure data obtained by the Knudsen effusion method. Unfortunately the precision of the data from which the thermodynamic values are calculated is insufficient to establish unequivocally a departure from the normal trend. The results are reported primarily because they substantiate the only previously reported activity measurements3 and because they include values for pure Cd which in contrast to those of the alloys show excellent precision. The Cd-Mg system is known to contain several ordered phases. However, this investigation was made at temperatures considerably above the transition temperatures, and the alloys may be regarded as nearly random solid solutions. EXPERIMENTAL Apparatus—The effusion cells are machined from solid tantalum rod. The lids consist of threaded, flanged rings of tantalum onto which are welded shallow cups of 4 mil molybdenum foil. The orifice, which is located at its center, is made by grinding off a small dimple produced with a sharpened punch. Such an orifice has been mounted in plastic, sectioned, and the thickness of the orifice edge measured in a microscope. The maximum thickness is 0.00016 in., and the Clausing factor is consequently equated to unity. The area of the orifice is determined by tracing its projection on the screen of the metallograph at a known magnification and measuring the area of the tracing with a planimeter. Excellent temperature control is achieved with a 4 by 18-in. alundum core, gradient-wound, furnace

Jan 1, 1960

By W. L. Eaton, S. H. Ash

A barrier pillar in the anthracite region of Pennsylvania is a portion of the bed that is left unmined along the property line or lines of adjoining mining properties, or between mines, or between parts of mines. Barrier pillars perform important functions in anthracite-mining operations. Their principal function is to act as dams to prevent water that accumulates in a mine from suddenly breaking into an adjacent mine or into adjacent mine workings and causing loss of life, property, or both. More than 83 billion gal of water is impounded in II; underground water pools in anthracite mines.1,2 These pools are confined by a system of barrier pillars which separate adjacent mines. The principal factor that threatens to cut short the life of the anthracite industry, to curtail production, and to affect the economic structure of the people and business dependent on anthracite for their livelihood is the inundation of anthracite mines.l,2,4,13,29 To obtain a clear picture of the anthracite mine-water problem, it is necessary to understand the part played by the barrier pillars that separate adjoining mines. Purpose of Report The purpose of this report is to show the methods of determining the size of barrier pillars in the anthracite region, to present the hazards that exist because of the lack of sound engineering being applied to the size of many existent barrier pillars, and to suggest a method whereby barrier pillars can perform the function for which they are intended. The Barrier-pillar Problem Article 111, section '10, of the Anthracite Mining Laws of Pennsylvania reads thus: It shall be obligatory on the owners of adjoining coal properties to leave, or cause to be left, a pillar of coal in each seam or vein of coal worked by them, along the line of adjoining property, of such width that, taken in connection with the pillar to be left by the adjoining property owner, will be a sufficient barrier for the safety of the employees of either mine in case the other should be abandoned and allowed to fill with water; such width of pillar to be determined by the engineers of the adjoining property owners together with the inspector of the district in which the mine is situated, and the surveys of the face of the workings along such pillar shall be made in duplicate and must practically agree. A copy of such duplicate surveys, certified to, must be filed with the owners of the adjoining properties and with the inspector of the district in which the mine or property is situated. Mining practices with regard to barrier pillars have varied greatly since mining began. In many anthracite mines the barrier pillars have been weakened or punctured. As a consequence, the condition

Jan 1, 1949

By Haydn H. Murray

The term clay is somewhat ambiguous unless specifically defined because it is used both as a rock term and as a particle size term. Actually most persons using the term clay realize that it has a double meaning and in most instances define it. As a rock term, clay is difficult to define because of the wide variety of materials that compose it so that the definition has to be general. Clay is a natural, earthy, fine-grained material composed largely of a limited group of crystalline minerals known as the clay minerals. Many definitions state that a clay has the property of plasticity when wet, but some clays are not plastic, i.e., halloysite and flint clay; however, most clay materials are plastic when wet. As a particle size term, clay is used for the category that includes the smallest particles. The maximum size particles in the clay size grade are defined differently on some grade scales. Soil investigators and mineralogists generally use 2 microns as the maximum size. The Wentworth scale' defines clay as material finer than approximately 4 microns. Grim' has found it convenient to use the term clay material for any fine-grained, natural, earthy, argillaceous material and in this way can include clays, shales, or argillites, and some soils if they are argillaceous. Even though no standard definition of the term clay is accepted by geologists, agronomists, engineers, and others, still the term is generally understood by all who use it. Clay is an abundant natural raw material and has an amazing variety of uses and properties which will be discussed in this chapter. In order to understand better the relationships between the clays and their uses, a brief discussion of the structure, chemical composition, and identification of the clay minerals is given. Mineralogical Composition of Clays Clays are composed of certain groups of hydrous aluminum, magnesium, and iron silicates that may contain calcium, potassium, and other ions, which are called the clay minerals, and other minerals of variable composition, which are nonclay minerals. Both clay minerals and nonclay minerals influence the properties of a clay material. In order to understand the relationship between clay minerals and the properties of a clay material, a short summary of the structure and composition of clay minerals is necessary. Any scheme of classification tends to impose a certain rigidity and as new and additional data become available it is necessary to revise classifications. Brindley3 has discussed the classification and structural mineralogy of clays in very lucid fashion and chose a classification based on structural features and chemical composition. Grim2 proposed a tentative classification which is based on the structural attributes of the various clay minerals. This classification is relatively simple and workable, and is shown as follows: I. Amorphous Allophane group II. Crystalline A. Two-layer types (sheet structures composed of units of one layer of silica tetrahedrons and one layer of alumina octahedrons). 1. Equidimensional Kaolin group kaolinite, dickite, nacrite 2. Elongate Halloysite group

Jan 1, 1960

By Thomas T. Read

MINING, which is at least twenty centuries old, was at first, and long, wholly a practical art. Little more than two centuries have elapsed since the inception of the idea that general education and a knowledge of natural science would be of advantage to those engaged in the mining industry. Though schools of mining began to develop in Europe early in the eighteenth century the concept of educational preparation for technical conduct of the industry was slow in winning general acceptance in the industry itself. John Hays Hammond, in his autobiography,1 recalls that when he sought employment with a mining company after he had graduated from Yale and Freiberg he had to assure its executive, before being even considered for employment, that he did not consider that he had learned anything about mining. This was as late as 1880. Existence in the industry of a prejudice against education of the academic type is perhaps an explanation why inadequate provision has been made by the mining industry for the continuing education of technical employees after they enter employment, although engineering education prior to engaging in industry has now won general acceptance. I am familiar with the existence of "loop" courses for engineering graduates in metallurgical plants, of schemes for giving mining graduates "practical experience," and of the interest in vocational courses for workmen which sprang up in the period of labor shortage during and immediately following World War I (but which quickly died when the depression of 1921 again made experienced workmen available in sufficient supply), having been chairman of a committee appointed by the Mineral Industry Education Division to make a survey of existing provisions for post-employment training for engineers in our industry. The results of that survey, reported elsewhere, clearly indicate that it is inadequate as to amount; the aim of this paper is to present evidence for belief that it is also inadequate in quality. To avoid becoming too involved, no reference will be made here to post-employment vocational training for workmen, which is a subject differing so much from post-employment education for engineers as to require separate treatment. Neither will postgraduate education for metallurgical engineers be discussed, because a distinction must then be made between training for research work and for plant operation, whereas mining engineers engage almost exclusively in operations. Since the situation in the coal-mining industry has been discussed at several recent meetings, that also will be passed over, and what follows should be taken as a discussion applying exclusively to the metal-mining industry. PROFESSIONAL DEVELOPMENT In the operations field (both in mining and metallurgy) the work to which a beginner is assigned is seldom closely related to previous academic education. The contrast is almost as great as between

Jan 1, 1941

By C. E. Shoenfelt

The most important development in an otherwise colorless year in Colorado oil fields was the completion, late in 1933, of a well on Iles dome on the western slope of Colorado, with an initial production of 1100 bbl. of oil. No great amount of activity beyond the drilling of necessary offset wells is to be expected as a result of the completion of a well of this size because the acreage on Iles dome is closely held; there is no pipe line outlet other than to the Texas Company's small refinery at Craig, 24 miles distant; and the market for the oil is limited to the demands of sparsely settled communities in the high mountains. The Iles field is located on a large dome developed in Mancos shale of Cretaceous age, and having a closure of approximately 400 ft. The structure is highly faulted on the southern end. The first commercial oil well on this dome was completed by the Midwest Refining Co. on the south side of sec. 22-4N.-92W. Oil of 37" gravity was developed in Nancos shale at 2560 ft. and water in the Dakota sandstone at 2630 ft., but a considerable amount of additional oil was found in the shale immediately overlying the Dakota. From 2630 ft., the well produced in 16 hr. 750 bbl. of fluid, 33 per cent of which was water. Production in the Sundance sand of Jurassic age was found in the Midwest Refining Company's No. 4 Parkinson, in the NE.-SW.-SE. of sec. 22-4N.-92W., and completed on the last day of the year 1926, for an initial production of 450 bbl. of oil at a total depth of 3242 ft. The well completed in 1933 is also a Sundance sand producer and was finished at a depth of 3447 ft. It extends the productive area of the field about a quarter of a mile cast. The Iles field is said to have a potential production of 5000 bbl. of oil a day from its present wells, but for the past year has been held at about 600 bbl. a day. The most important test, stratigraphically, ever drilled on the western slope of Colorado is the California Company's No. 1 Raven, NW.-SE. of sec. 30-2N.-102W., which had an initial production of 138 bbl. of oil of 32" gravity, in the Weber formation (Pennsylvanian) between 5750 to 6315 ft. It mas drilled to 7137 ft., probably near the base of the Weber, and then plugged back to 6315 ft. It was shot quite heavily with nitro-glycerin, but failed to respond. While the well can hardly be classified

Jan 1, 1934

By Thomas T. Read

MINING, which is at least twenty centuries old, was at first, and long, wholly a practical art. Little more than two centuries have elapsed since the inception of the idea that general education and a knowledge of natural science would be of advantage to those engaged in the mining industry. Though schools of mining began to develop in Europe early in the eighteenth century the concept of educational preparation for technical conduct of the industry was slow in winning general acceptance in the industry itself. John Hays Hammond, in his autobiography,1 recalls that when he sought employment with a mining company after he had graduated from Yale and Freiberg he had to assure its executive, before being even considered for employment, that he did not consider that he had learned anything about mining. This was as late as 1880. Existence in the industry of a prejudice against education of the academic type is perhaps an explanation why inadequate provision has been made by the mining industry for the continuing education of technical employees after they enter employment, although engineering education prior to engaging in industry has now won general acceptance. I am familiar with the existence of "loop" courses for engineering graduates in metallurgical plants, of schemes for giving mining graduates "practical experience," and of the interest in vocational courses for workmen which sprang up in the period of labor shortage during and immediately following World War I (but which quickly died when the depression of 1921 again made experienced workmen available in sufficient supply), having been chairman of a committee appointed by the Mineral Industry Education Division to make a survey of existing provisions for post-employment training for engineers in our industry. The results of that survey, reported elsewhere, clearly indicate that it is inadequate as to amount; the aim of this paper is to present evidence for belief that it is also inadequate in quality. To avoid becoming too involved, no reference will be made here to post-employment vocational training for workmen, which is a subject differing so much from post-employment education for engineers as to require separate treatment. Neither will postgraduate education for metallurgical engineers be discussed, because a distinction must then be made between training for research work and for plant operation, whereas mining engineers engage almost exclusively in operations. Since the situation in the coal-mining industry has been discussed at several recent meetings, that also will be passed over, and what follows should be taken as a discussion applying exclusively to the metal-mining industry. PROFESSIONAL DEVELOPMENT In the operations field (both in mining and metallurgy) the work to which a beginner is assigned is seldom closely related to previous academic education. The contrast is almost as great as between

Jan 1, 1941

By George D. Roberts

Civil engineers are primarily interested in maximum core recovery. This is even more important in foundation work than in mining investigations where sludge samples are of some value. The soft materials, such as clay, gouge, or partially decomposed seams that are most readily ground up or washed away, are the materials that usually have the greatest influence on foundation and leakage conditions. Undisturbed samples of such materials should be obtained for detailed geological examination and laboratory tests. Therefore the attainable goal of sampling in connection with foundation investigations is the recovery of 99.9 pot of undisturbed core. Factors Causing Core Losses The principal factors contributing to core losses are: (1) coring badly fractured or very soft material; (2) Use of improperly designed coring equipment, including the use of bits of too small a diameter; (3) lack of the proper training and attitude on the part of the drillers and pertinent technical information on the part of the supervising personnel. Little can be done about the character of the material being cored but the other two factors contributing to unsatisfactory performance can be corrected. Although the title of this Paper is the improvement and standardization of core barrels, it is believed that the drillers' shortcomings and the lack of pertinent technical literature are more detrimental to good core recovery than either of the other factors. Uses For Large Diameter Core Barrels It is well known that the use of larger diameter bits and core barrels generally produce better cores, particularly when sampling soft or broken material. The use of larger and larger bit diameters and the insistence on better core recovery is well demonstrated by the change in practices of the various districts of the Corps of Engineers. Today NX core is the smallest diameter commonly permitted for investi-gational work and the more active districts are using coring bits up to 6 in. in diameter, as well as 36 and 40-in. calyx rigs. The latter size has proved to be just as economical because it provides more room for ctured removal and cleanup operations. The use of large diameter cores to replace the driving and jacking methods of obtaining soil samples is also increasing. The 57/8-in. Denison sampler, developed by H. L. Johnson to obtain samples from soft formations, is one example. The principal uses for bits larger than NX are: sampling soils and semi-consolidated rock; sampling decomposed rock or rock with soft seams; sampling concrete; obtaining samples for laboratory tests; drilling drainage holes, and other miscellaneous work. Core Drilling Practices of Several Engineer Districts Little Rock District The writer was employed in the Little Rock District from 1938 until 1942. Prior

Jan 1, 1949

By J. L. Taylor, P. Duwez

PARTIAL phase diagrams of titanium with iron, cobalt, and nickel have been established by previous investigators.1-3 These diagrams seem to be reliable, at least for concentrations of titanium ranging from 0 to about 50 at. pct. More recently, Long and his collaborators have presented a tentative titanium-nickel diagram covering the titanium-rich side.' These various studies indicate that in each system three intermediate phases exist and correspond to the compositions Tia, TiX, and TiX2 (or TiX3). The purpose of this paper is to review the available information on the crystal structure of these phases and to give the results of additional X-ray diffraction studies. Previous Work A summary of the work of Wallbaum and his collaborators"-' is presented in table I. The first intermediate phase on the titanium-rich side of the three systems, with iron, cobalt, and nickel, is of the Ti2X type and has a face-centered cubic structure. Although it is stated that the unit cell contains 96 atoms: no lattice parameters are given. The second phase is centered around the compositions TiX for the three systems. The crystal structure of this phase is mentioned in a short note" as being body-centered cubic (CsCl type), but no parameters are given. The analogy between the three systems, which is perfect as far as the two phases of the type Ti2X and TiX are concerned, disappears at the titanium-poor end of the diagrams. In the system titanium- iron, TiFe2 has an hexagonal (MgZn2 type) struc-ture with the parameters given in table I. The existence of TiFe3 has also been reported8 (table I) but was questioned by other investigators2 and is also refuted by the present study. In the system titanium-cobalt, a phase corre-sponding to the composition TiCo2 has been investi-gated by Wallbaum and Witte6 who claim that the structure is cubic (MgCu2 type) for alloys slightly deficient in cobalt and hexagonal (MgNi2 type) for alloys slightly deficient in titanium. The parameters of the two phases are given in table I. In the system titanium-nickel, the nickel-rich phase corresponds to the composition TiNi3 and has been described as hexagonal7 with the parameters given in table I. The present investigation was initiated for the purpose of filling in the gaps in the knowledge con-cerning the structure of the various phases sum-marized in table I. Since the titanium used by Wallbaum and his collaborators was only 95 pct pure,' it was found desirable to redetermine the lattice parameters of the alloys prepared with a

Jan 1, 1951

By A. R. Kinkel

THE Shasta copper-zinc district of northern California lies in the foothills of the Klamath Mountains at the northern end of the Sacramento Valley. It contains two main areas of base-metal ore deposits, the East Shasta and West Shasta districts shown in Fig. 1. The East Shasta includes the area from the Afterthought mine, near the settlement of Ingot on U. S. highway 299 East, to the Bully Hill mine, on the north side of the Pit River, 9 miles to the northwest. The West Shasta, a well-defined northeast-trending district about 8 miles long and 2 miles wide, west of the Sacramento River, Fig. 2, is the western part of the ore-bearing area formerly known as the Shasta copper belt. Referred to in the older writings as a copper arc, this belt was thought to form a crescent extending entirely around the head of the Sacramento Valley, the convex side to the north. Recent studies by the U. S. Geological Survey in cooperation with the California State Division of Mines show that there are two districts, one at each end of the so-called copper arc, which are distinct in geologic structure and ore occurrences, and that there is only sporadic and unconnected copper mineralization between them. Ore deposits of the two districts are similar in mineralogy, that is, they are largely massive pyritic ores containing copper and zinc with minor amounts of gold and silver, but the structure and structural control of ore deposition are different. The ores of the East Shasta district are replacements of steeply dipping shear zones in Triassic rhyolite and shale, whereas those of the West Shasta district are replacements of gently dipping rhyolite of Middle Devonian age and are essentially flat-lying. Thus there is marked lateral control of orebodies by shear zones in the East Shasta district and equally marked vertical control of orebodies by the stratigraphy in the West Shasta district. Geologic Setting The Paleozoic rocks of the West Shasta district range in age from Middle Devonian (?) to Missis-sippian. The oldest formation exposed is the Copley greenstone, probably Middle Devonian. It is composed of volcanic flows, volcanic breccias, and tuffs, all of mafic composition, but contains minor amounts of shale and rhyolitic tuff. Largely submarine in origin, it rests upon a base of older marine sediments, as shown diagrammatically in Fig. 3. The lower part of the formation is made up principally of flows, many of which are ellipsoidal, and the upper part contains most of the pyroclastic material. The formation is at least 3700 ft thick. The Balaklala rhyolite overlying the Copley greenstone is made up of many rhyolitic flows and beds

Jan 1, 1956

By Harry H. Angst

THE history and efficiency of 40-ton capacity dump trucks for surface waste removal at the New Cornelia opencut copper mine, at Ajo, Ariz., are summed up in this paper. Tabulations of truck performance, together with detailed cost figures, prove the value of truck transportation for certain special operations at a large mine. Emphasis is given to the fact that truck transportation not only reduces the haulage costs but also reduces the breaking ground and shovel-loading costs per ton. The New Cornelia opencut copper mine is in the desert. The climate is ideal for seven months of the year, but the five summer months are uncomfortably hot. The elevation is 1820 ft. above sea level and the annual rainfall is approximately 6 inches. The ore body is a monzonite-porphyry deposit. The waste-rock capping consists of hard rhyolite, conglomerate and porphyry. All the material is hard and cannot be excavated without blasting. The rhyolite and conglomerate waste is very blocky. The weight of the rock in place is 2.16 tons per cubic yard. The average mine production per day is 22,000 tons of ore and 18,000 tons of waste rock, the latter consisting of surface rock and material within the ore body too low in grade to send to the concentrator. The rock is loaded by electric shovels with 4 ½ -cu. yd. dippers. The primary transportation of ore to the concentrator and waste to the dumps is by 30-yd. dump cars and oil-burning steam locomotives; therefore truck haulage of waste is a minor factor of the New Cornelia operation. However, as New Cornelia was one of the first mines to use large-capacity dump trucks for haulage, the results obtained are of interest. EQUIPMENT On the south side of the New Cornelia is a hill rising about 500 ft. above the normal surface elevation, roughly 1000 ft. in diameter at its base, called locally Arkansas Mountain peak. Exploratory drilling in this area developed additional ore that could be made accessible for open pit operation only by moving the mountain face a maximum distance of 400 ft. to the south. The slope of the mountain adjacent to the pit averaged 35°, and the small area involved precluded the use of train haulage for the removal of the 4,000,000 tons of waste estimated above the normal pit elevation. A study of dump-truck haulage at Boulder Dam and San Gabriel Dam led to the decision to use large-capacity end-dump trucks rather than smaller trucks, for the following reasons: 1. The truck would have to withstand the impact of hard, heavy chunks, some of them weighing as much as 10 tons, which would be dumped from the 4 ½ -cu. yd. dipper. 2. The coarse, rocky, abrasive material would soon damage a light truck body. 3. The body of a small truck could not be efficiently loaded with coarse rock.

Jan 1, 1941

By Henry H. Angst

Tim history and efficiency of 40-ton capacity dump trucks for surface waste removal at the New Cornelia opencut copper mine, at Ajo, Ariz., are summed up in this paper. Tabulations of truck performance, together with detailed cost figures, prove the value of truck transportation for certain special operations at a large mine. Emphasis is given to the fact that truck transportation not only reduces the haulage costs but also reduces the breaking ground and shovel-loading costs per ton. The New Cornelia opencut copper mine is in the desert. The climate is ideal for seven months of the year, but the five summer months are uncomfortably hot. The elevation is 1820 ft. above sea level and the annual rainfall is approximately 6 inches. The ore body is a monzonite-porphyry deposit. The waste-rock capping consists of hard rhyolite, conglomerate and porphyry. All the material is hard and cannot be excavated without blasting. The rhyolite and conglomerate waste is very blocky. The weight of the rock in place is 2.16 tons per cubic yard. The average mine production per day is 22,000 tons of ore and 18,000 tons of waste rock, the latter consisting of surface rock and material within the ore body too low in grade to send to the concentrator. The rock is loaded by electric shovels with 4 1/2-cu. yd. dippers. The primary transportation of ore to the concentrator and waste to the dumps is by 30-yd. dump cars and oil-burning steam locomotives; therefore truck haulage of waste is a minor factor of the New Cornelia operation. However, as New Cornelia was one of the first mines to use large-capacity dump trucks for haulage, the results obtained are of interest. EQUIPMENT On the south side of the New Cornelia is a hill rising about 500 ft. above the normal surface elevation, roughly 1000 ft. in diameter at its base, called locally Arkansas Mountain peak. Exploratory drilling in this area developed additional ore that could be made accessible for open pit operation only by moving the mountain face a maximum distance of 400 ft. to the south. The slope of the mountain adjacent to the pit averaged 35°, and the small area involved precluded the use of train haulage for the removal of the 4,000,000 tons of waste estimated above the normal pit elevation. A study of dump-truck haulage at Boulder Dam and San Gabriel Dam led to the decision to use large-capacity end-dump trucks rather than smaller trucks, for the following reasons: I. The truck would have to withstand the impact of hard, heavy chunks, some of them weighing as much as 10 tons, which would be dumped from the 4 ½ -cu, yd. dipper. 2. The coarse, rocky, abrasive material would soon damage a light truck body. 3. The body of a small truck could not be efficiently loaded with coarse rock.

Jan 1, 1941

By W. J. West, T. D. Mueller, J. E. Warren

Two methods are used to obtain the relatiorlship between ratio of gas to oil permeability and gas saturation. One method is to calculate a curve from field measurements of reservoir behavior: the other is to measure the relationship on a core .sample in the laboratory. The curves resulting from these TWO methods arc often comparer1 and seldom found to agree. The most important reason for the nonagreement is often overlooked. This reason is that the performance k9/k0 curve is calculated with the assumption of uniform pressure and saturution throughout the reservoir. To show the differerence between performance type kg/k0 curves and laboratory type kg/k0 curves. two performance type kg/k0 curves for two different production rates were calculaled from the production perforrnance of a hypothetical reservoir for which a laboratory type curve was assumed. Production performance of the hypothetical reservoir bras calcu- lated by use of international Business Machine's 701 Computer. The perfortnance type kg/k0 curves differed from the laboratory curve for the hypothetical reservoir. The performance type crrrve for the lowest production rate most nearly coincided with the laboratory curve. INTRODUCTION Predictions of oil reservoir performance are often made with the aid of complex numerical calculations. An important quantity required in these calculations is the ratio of gas permeability to oil permeability (kg/k0 )at a particular gas saturation. Two methods arc used to obtain a relationship between this ratio and gas saturation. The curves resulting from these two methods are called performance kg/k0 curves and laboratory kg/k0 curves. The performance curve is obtained from field measurements of reservoir behavior. Particular kg/k0 values are calculated from the producing gas-oil ratio, the cumulative reservoir withdrawals, the original oil in place, the average reservoir pressure, and the PVT properties of the reservoir fluids. The particular kg/k0 values are then plotted against the average rcservoir gas saturations for the entire history period. In order to use this derived kg/k0 relationship for performance predictions, a curve through these points is extrapolated to higher gas saturations. The laboratory kg/k0 curve is obtained from measurements made on core samples. Saturation is maintained uniform throughout the core during these measurements. The kg/k0 curves obtained with these two independent methods are often compared and seldom are found to be the same. Various reasons have been offered for the lack of correlation. One is that insufficient core? were obtained to derive a curve representative of the reservoir rock. Another is that laboratory methods have failed to reproduce reservoir conditions. The most important reason for the noncorrelation has been often overlooked because of the difficulty in evaluating its effect. This reason is that the performance or field kg/k0 curve is calculated with the assumption of uniform pressure and saturation throughout the reservoir. In any

Jan 1, 1956