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By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

By C. E. McCarthy, A. George Stern, Stanley J. Green

The national interest prompts consideration of any new source of mineral wealth even though the immediate need may be of minor importance. A critical shortage of potash in the United States during the World War of 1914-18 stimulated broad development, which established an American potash industry on a basis so sound that this commodity is not even listed as urgent in the present war program. However, this should not minimize its importance, and the need remains, not only for adequate supplies of the usual potash salts (the chloride and the sulphate) but of all potash chemicals. Potassium carbonate (known commercially as pearl ash) has been a high-cost commodity; in consequence, the quantity used by industry has been relatively small. By a combination of fortuitous geographical circumstances, there is opportunity to produce a low-cost potassium carbonate with the possibility that this might become an important source of chemical potash. The following paper considers a recently developed process1 taking advantage of these conditions. Wyomingite Deposit For many years it has been known that a deposit of potash-bearing igneous rock known as wyomingite occurs in the south- western part of Wyoming at Zirkel Mesa, which is 44 miles northeast of the town of Green River (Fig. I). The western escarpment of this deposit is about 3 miles east of the Superior Branch of the Union Pacific Railroad. This mesa alone covers approximately 3300 acres, and the average thickness of the wyomingite flow is 75 ft. It has been estimated2 that about one billion tons of wyomingite occur in this one deposit, all of which can be mined by open quarrying without overburden. In addition, there are several smaller deposits within 10 miles, so that the aggregate wyomingite in the area approaches two billion tons. This is one of the world's largest sources of leucite, and the only known deposit of its kind in the United States. The wyomingite (which is in the form of a lava flow and numerous small craters), with the mineral leucite as its most important component, has a mineral composition3 of: leucite, 50 per cent; phlogopite, 15; diopside, 12; kataphorite, 15; glass and apatite, 8. The phlogopite is present in visible phenocrysts 1/8 to 1/2 in. in diameter, but the leucite usually is almost microcrystalline and not capable of liberation by grinding. Chemically, the Wyomingite has a potash Content averaging slightly higher than eleven per cent K2O. The potash is mainly in the leucite, but part of it is in the phlogopite and a small amount is in the feldspars that are occasionally seen as undigested occlusions of the granite country rock, through which the leucite-bearing lava intruded, and as sanidine feldspar both in the crystalline and glassy phases of the rock. A typical analysis of the wyomingite from Zirkel Mesa is as follows:

Jan 1, 1948

By R. Stefanko

The use of strain gauges and compression pads for an underground bolt installation in a central Pennsylvania coal mine revealed that satisfactory anchorage could not be maintained in the shale stratum that constituted the immediate roof. Comprehensive laboratory tests resulted in a new shell designed with an increased area of contact and smaller surface serrations. Retightening of the bolt was also found to be beneficial for maintenance of high load values. Roof bolting in coal mines has progressed very rapidly in the last decade, and today this method of support is in general use. There are many theories concerning the functions of roof bolts, but the importance of beam building is indisputable. Proper tension in the bolts minimizes shearing along bedding planes and creates a thick, competent beam from individually weak layers. Bolting must be accomplished immediately upon exposure of the roof, and bolt tension must be maintained if roof failure is to be prevented. Seams are being mined today with roof bolt support where conventional timbering was not effective.' The widespread use of continuous mining equipment was made possible by roof bolts which permit long supported spans free of obstructions. Since adequate anchorage is the key to successful bolting, better methods for determining the efficiency of anchorage are continually being sought. Attempts have been made to measure bolt loads, thus assessing anchorage efficiency by employing sensing devices between the bolt head and roof. 3-s None of the techniques appears to be very accurate, and all have at least one serious deficiency. The greatest efforts are being directed toward the standardization of the pull test.6" Nevertheless, the author believes that the pull test is only one criterion for establishing the efficiency of anchorage and not necessarily the most important one. The conventional pull test involves the short-term application of forces, whereas the success of roof bolt installations is dependent upon the ability of anchorage to sustain relatively long-term forces. Therefore, it would appear that the time-dependent function of the anchor is the single most important parameter for the success of roof bolting in many types of rocks. UNDERGROUND TESTS Strain gauges and compression pads were employed to measure bolt loads in a central Pennsylvania coal mine.8 Bolting had been employed only sporadically in the shale top since dubious benefits were obtained with this type of support. Table I shows some typical results from this investigation. Anchorage losses were extremely variable among the bolts, but all showed a considerable degree of bleed-off. The tests indicated that anchorage efficiency was extremely low in this type of shale. The results confirmed that bolting under these conditions was practically worthless, and thus a waste of material and labor. Probably even more important, bolting constituted a hazard when reliance was placed upon it for support. Certainly the complete lack of confidence in roof bolting, which was apparent in the mine personnel, was justifiable since the bolts did not accomplish their purpose. PRELIMINARY LABORATORY BOLT TESTS The extreme losses in bolt tension that were detected in underground installations prompted a comprehensive study to assess the causes. However, preliminary studies were first made to determine the combined stresses in a roof bolt and to prove conclusively that the observed bleed-off phenomena were due to anchorage: slippage and not to some other physical manifestation. In all of the ensuing tests

Jan 1, 1962

By A. George Stern, Stanley J. Green, C. E. McCarthy

The national interest prompts consideration of any new source of mineral wealth even though the immediate need may be of minor importance. A critical shortage of potash in the United States during the World War of 1914-18 stimulated broad development, which established an American potash industry on a basis so sound that this commodity is not even listed as urgent in the present war program. However, this should not minimize its importance, and the need remains, not only for adequate supplies of the usual potash salts (the chloride and the sulphate) but of all potash chemicals. Potassium carbonate (known commercially as pearl ash) has been a high-cost commodity; in consequence, the quantity used by industry has been relatively small. By a combination of fortuitous geographical circumstances, there is opportunity to produce a low-cost potassium carbonate with the possibility that this might become an important source of chemical potash. The following paper considers a recently developed process1 taking advantage of these conditions. Wyomingite Deposit For many years it has been known that a deposit of potash-bearing igneous rock known as wyomingite occurs in the south- western part of Wyoming at Zirkel Mesa, which is 44 miles northeast of the town of Green River (Fig. I). The western escarpment of this deposit is about 3 miles east of the Superior Branch of the Union Pacific Railroad. This mesa alone covers approximately 3300 acres, and the average thickness of the wyomingite flow is 75 ft. It has been estimated2 that about one billion tons of wyomingite occur in this one deposit, all of which can be mined by open quarrying without overburden. In addition, there are several smaller deposits within 10 miles, so that the aggregate wyomingite in the area approaches two billion tons. This is one of the world's largest sources of leucite, and the only known deposit of its kind in the United States. The wyomingite (which is in the form of a lava flow and numerous small craters), with the mineral leucite as its most important component, has a mineral composition3 of: leucite, 50 per cent; phlogopite, 15; diopside, 12; kataphorite, 15; glass and apatite, 8. The phlogopite is present in visible phenocrysts 1/8 to 1/2 in. in diameter, but the leucite usually is almost microcrystalline and not capable of liberation by grinding. Chemically, the Wyomingite has a potash Content averaging slightly higher than eleven per cent K2O. The potash is mainly in the leucite, but part of it is in the phlogopite and a small amount is in the feldspars that are occasionally seen as undigested occlusions of the granite country rock, through which the leucite-bearing lava intruded, and as sanidine feldspar both in the crystalline and glassy phases of the rock. A typical analysis of the wyomingite from Zirkel Mesa is as follows:

Jan 1, 1948

By Tell Ertl

THE Piceance Creek Basin in northwestern Colorado is believed to contain the richest large deposit of oil shale in North America. The major portion, about 1650 sq miles, is bounded by the White River on the north, the Grand Hogback, which runs from Meeker to Rifle, on the east, the Colorado River to the south, and the drainage basin of Salt and Douglas Creeks on the west. The Basin is a plateau 7000 to 9000 ft high, composed of 3000 ft of the Green River series of Eocene age and surrounded by a steep escarpment of 1000 to 3000 ft. From the escarpment the beds dip gently toward the center. Although oil shale is found in all members of the Green River series, the rich thick beds are found chiefly in the Parachute Creek member. This is normally capped by 100 to 200 ft of the Evacuation Creek, the youngest member of the Green River series. Actually the Piceance Creek Basin oil shale is not a shale, nor does it contain oil. It is magnesium marlstone holding solid organic matter capable of being broken down by heat into gases, some of which can be condensed into oil. The marlstone is resilient nonporous rock, difficult to sample by pick or hammer and moil. Many companies and individuals prospected the Colorado oil shale shortly after the first World War,1 see Fig. 1, but because of the discovery of large quantities of petroleum in the late 1920's in California, Oklahoma, and East Texas, interest in oil shale development waned and did not revive until World War II. In 1944 the writer, employed by the Bureau of Mines, prospected the high grade oil shale in numerous places along the southern outcrop of the Piceance Creek Basin. Several channel samples about 100 ft in length were cut, see Fig. 2. An eightman crew needed a week to cut a satisfactory channel 80 ft long in the cliff face. The Phillips Petroleum Co., the Union Oil Co. of California and others also prospected the oil shale outcrop. In 1945 the Bureau of Mines initiated a core drilling program for the purpose of obtaining accurate samples from the Piceance Creek Basin. The U. S. Geological Survey set up a program for the geological study of the Basin that is still going on.2 Until the end of 1951 a total of 31 holes was drilled for core with diamond drills, and, in addition, cuttings through the oil shale have been obtained from 4 gas wells drilled by standard rotary oil well rigs. The Bureau of Mines drilled 12 of the holes, the Union Oil Company of California 8, the Standard Oil Company of California 6, the Sun Oil Company 3,

Jan 1, 1952

By Roland Kammel

For electrolytic recovery of silver from dilute aqueous solutions improved mass transfer is necessary to achieve favourable current efficiencies and high space-time yields. The many cell designs proposed for this application can be divided into two different groups: cells from which the silver must be redissolved and cells from which the deposit can be reclaimed as a metal. Introduction Electrolytic silver recovery methods have proven advantageous under technical and economic aspects for the treatment of rinse water, still rinses and spent silver solutions in the electroplating and surface finishing industry, of exhausted fixing baths from photographic processing, bleed electrolytes from silver refining processes and leach liquors. Examples of the compositions of typical dilute silver solutions are listed in Table I.

Jan 1, 1984

By J. C. Kerkering, P. M. Daling

This paper discusses a study performed to evaluate the feasibility of transferring formal system safety assessment technology to the mining industry. A representative listing of formal techniques was evaluated to select the most useful methods for the mining industry to implement. Criteria for selecting methods included the interests, needs, and requirements of mine safety officials. This paper describes one of the selected methods, preliminary hazards analysis (PHA), and illustrates the analysis procedure for a mining example. This study also includes an assessment of the adaptability of selected methods to a user-interactive computer system for use by mine safety personnel. This paper includes a description of a conceptual computerized PHA-type safety program that facilitates performance of the safety analysis and can accommodate updates and revisions as necessary.

Jan 1, 1983

By Georges Caraghiaur

INTRODUCTION A small, independent mining enterprise (SIME) is an independently owned and operated business concern engaged primarily in mineral development, from the exploration to the production stage. By definition, a SIME is not dominant in its field and has no long-term financial backing. It is not a major producer of minerals, nor is it a petroleum company, a government organization or a subsidiary of such organizations. A SIME may or may not be listed on stock exchanges. The SIME is also called small miner, junior mining company, speculative mining company, and penny-stock mining company. The survival and growth of a SIME is most often controlled by an independently minded entrepreneur, who not only conceives and carries out new projects, but is materially affected by the success or failure of these projects. Usually, a SIME does not have a large or constant cash flow. Such a lack of long-term financial backing forces the SIME to meet the challenge of finding capital for its activities on a step-by-step basis. Its survival strategy is thus based on finding funds at the right time, of the right type and in the right quantities. The purpose of this paper is to show that the type of venture in which a SIME is involved affects the difficulty of raising new capital and the choice of possible sources of funds. The Small, Independent Mining Enterprise (SIME) The SIME is not a figment of the imagination of financial speculators. In Canada, there are more than 1,800 small mineral enterprises which spend on the average less than $1,000,000 per year on activity which is directed at bringing forth new mineral supply. In the United States, 64% of the non-fuel mining operations are conducted by nine or fewer men, and the coal industry is characterized by several thousand smaller independent producers, still predominantly family owned or controlled. In Chile, 80% of the mines can be classified as small mines. In the Philippines, approximately 60% of metal mining operations are classified as small mines. The SIME is actively engaged in mining all over the world, and very much so in the Pacific Rim. The Mineral Supply Process The whole process of bringing forth new mineral supply involves a dynamic sequence of activities, ranging from the initial exploration for a mineral occurrence to the production of a marketable product. A typical sequence is outlined in Figure 1. At the grass-roots exploration stage, a series of geological, geophysical and geochemical tests is aimed at locating a mineral occurrence. Especially in the early phases of such activities, the uncertainty (risk) surrounding a venture is high, since only general geological information exists on which to base a judgement of the economic viability of the venture. Such risk is attractive to few investors. As the information-gathering process continues, uncertainty diminishes. Once exploration efforts result in the discovery of a mineral occurrence, i.e. when indications of mineralization and mineable widths are obtained by drilling, there is still uncertainty concerning the extent of mineralization and the probability of having located an economic mineral deposit. The expected return on investment at such a stage may well be negative but the slight possibility of a very high positive return will make a number of financial sources available to an exploration company. At the development stage, much of the uncertainty

Jan 1, 1985

By S. A. Hovanessian

Analytical solutions are obtained for calculating the pressure distribution in rectangular fields due to injection and/or producing wells located anywhere within the field. The field is assumed to be homogeneous with either constant pressure or no-flow boundaries. These solutions are extended to include the calculation of average reservoir pressure and permeability from pressure build-up curves. Numerical examples are given to illustrate the application of equations to practical problems. INTRODUCTION Pressure distribution analyses are of considerable importance in the field of reservoir mechanics. A number of papers dealing with this subject have appeared in the petroleum literature. perrinel in 1956 presented a summary of the methods available for pressure build-up calculations and discussed the applicability of each in some detail. More recently (1958), Hazebroek, et a1,2 discussed a theoretical method for obtaining pressure distribution in a radial field. Also in 1958, Nisle3 gave a theoretical solution of the pressure distribution in a field extending to infinity with a partially penetrating line source at the origin. Mathews, et a1,4 applied the method of images to existing solutions for pressure distribution in radial fields and obtained expressions for calculating pressures in bounded reservoirs. However, a literature survey revealed that a general analytical solution for the pressure distribution in a rectangular field, with injection and/or production wells located anywhere within the field, was not available. Such a solution could be used to approximate the pressure distribution in a field the shape of which is nearly rectangular. Furthermore, it could be applied to pressure calculations pertaining to a single well taken from a system of wells which were drilled in a rectangular pattern. In this case the boundaries of the well drainage area are considered to be the perpendicular bisectors of the lines joining the well and its nearest neighbors. The basic problem considered in this paper consists of a rectangular field with either constant-pressure 01 no-flow boundaries and with either an injection or production well located anywhere within the field. The physical properties of the rock and fluids present are considered to be constant. Analytical expressions are obtained for the pressure distribution within the field during the production or injection periods. These expressions, which are infinite series, can be used to evaluate pressure at any point within the field as a function of time and position. For fields containing more than one well (production and/or injection), the solutions for each well acting independently are superposed. That is, at any point in the field, the pressures due to each injection 01. production well acting alone can be algebraically added to give the pressure resulting from all wells acting simultaneously. Evaluation of the series solutions given in this paper is readily accomplished by means of an electronic digital computer; and, in most cases, sufficient accuracy is obtained by setting the upper limits of summation at 20. These solutions are extended to include the calculation of average reservoir pressure and permeability from build-up data. A number of numerical examples are given to illustrate application of the solutions to actual field problems. ANALYTICAL SOLUTIONS The partial differential equation representing the pressure distribution in a homogeneous rectangular field (see Fig. I.), having a single source or sink of strength (2, can be written*

By C. B. Scotty, E. F. Egan

INTRODUCTION The neutron-garnma log has been used for stratigraphic correlation by the oil industry for a number of years. In the past few years, the quantitative application of the log to provide information with respect to porosity has received considerable attention. The basic concept of quantitative interpretation has been pointed out in the literature.1 It shall be the purpose of this paper to develop by empirical means a method whereby the relationship of porosity to neu. tron-gamma deflection may be predicted for various bore hole diameters. This relationship is applicable to the present commercially available neutron-gamma log run in open hole. Data available at present are insufficient to establish similar correlations for cased holes. Through the development of such a correlation, the importance of the various factors which influence the log become evident. NEUTRON-GAMMA CURVE The logging tool conisists of a source of neutrons and a means of detecting and measuring gamma rays, the detector being shielded from the direct emission from the source. The gamma rays reaching the detector have three sources of origin: (1) the natural gamma radiation emitted by the formations, (211 induced gamma radiation resulting from neutron bombardment. (3) gamma rays emitted by the source and scattered by the formation. It has been found that a neutron derived porosity correlation is best for shale-free formations. Under this condition the natural gamma ray radiation is quite low, and therefore, will be neglected in this work. The induced gamma ray radiation resulting from neutron bombardment is predominantly a function of the hydrogen atom concentration of the substance penetrated. Due to the geometry of the presently used system, the neutron-gamma curve indicates low radioactivity opposite high hydrogen concentration and high radioactivity opposite low hydrogen concentration (Figs. 1 and 2). Having a curve which measures hydrogen atom concentration, it follows that if the matrix of the formation penetrated consists only of atoms other than hydrogen, and if the fluid filling the pores is either oil or water, the curve should be an indication of porosity (both oil and water have practically the same number of hydrogen atoms per unit volume'). The above criteria for a porosity (Correlation are fulfilled in clean sandztones and limestones in which the pore volume is completely filled with liquids (oil or water), but not in formations containing either free gas or shale or both. Free gas. even under pressure, contains considerably less hydrogen per unit volume than oil and water; therefore, a variation in response on the curve would be expected. Camp- bell and Winter,' indicate that this difference in hydrogen concentration is sufficient to distinguish between sands containing water or oil and those containing dry gas. Shale is a mineral that contains hydrogen in its molecular structure in addition to the hydrogen which is present in the water of hydration. Thus, in formations containing variable amounts of shale, the hydrogen concentration indicated by the curve is not a direct indication of porosity. Bush and Mardock' have indicated that a correction may be made by use of the gamma ray curve to provide a porosity correlation. Since the induced gamma rays are predominaiitly a function of hydrogen concentration, it would be anticipated that the fluid filled bore hole would materially influence the logs. It has been pointed out that the neutron-gamma curve response decreases with increase in hydrogen concentration, therefore, the response due to the induced gamma ray component decreases with increase in fluid filled hore hole diameter. The scattered gamma ray component of the log increases with increase in bore hole diameter. Although there is an

Jan 1, 1952

By W. D. Robertson

SINCE the comprehensive paper of Moore, Beckin-sale, and Mallinson,' little consideration has been given to the mechanism of mercury stress cracking of copper-base alloys, apart from extensive work on metallurgical and fabrication variables employing mercurous nitrate as an evaluation test. However, with the phase diagrams now available and the recent considerations of interfacial tension at grain boundaries, developed by Smith,2 it appears that the elements of a detailed solution to the problem are available. The facts that require explanation are: 1—Pure copper is not embrittled by mercury, with or without an applied or residual stress. 2—Cu-Zn and other copper-base alloys, in excess of some undefined solute concentration, are embrittled provided an external or residual stress is acting; cracking is not observed in the absence of stress. 3—The path of fracture is invariably intergranu-lar. 4—Copper and copper-base alloys do not dissolve readily in mercury at room temperature.' Consideration of the Cu-Hg phase diagram% hows that, at room temperature, solution of copper in mercury is probably limited by the formation on the surface of a microscopically thin layer of inter-metallic compound, CuHg, having a very limited range of solubility which effectively restricts diffusion of copper and mercury. But, above 150°C (the highest peritectic temperature) copper solid solution is in equilibrium with a liquid Hg-Cu solution and, consequently, copper should dissolve freely in a large volume of mercury with the formation of an equilibrium dihedral angle at grain boundaries in contact with the liquid phase. Depending on the magnitude of this angle, mercury will or will not penetrate the solid phase along grain boundaries and spread over grain faces. Thus the measurement of the dihedral angle of grain boundaries in equilibrium with the liquid phase above 150°C should indicate whether copper is intrinsically resistant to embrittlement by mercury. In view of the fact that mercury is appreciably soluble in copper, it may also be anticipated that diffusion of mercury will take place at higher temperatures in the absence of the intermetallic compound and possibly at a more rapid rate along grain boundaries than through the grains, resulting in a Cu-Hg alloy of unknown properties. In the case of brass, the ternary system Cu-Zn-Hg is involved and, from the resistance of brass to general solution in mercury at room temperature, it is probable that a compound of limited solubility range also exists, similar to the binary Cu-Hg sys- tem. Also, compared with copper, a change in the dihedral angle at grain boundaries in contact with mercury may be expected to result from the addition of zinc to copper. The preceding generalizations appear to be confirmed by the following experiments. Experimental Results Oxygen-free copper wire, annealed at 700°C for 3 hr in copper foil, was immersed in mercury at 170°C, with and without an applied stress. After 24 hr immersion, specimens which were spring loaded to a nominal stress of 10,000 psi had not failed, and there was no apparent embrittlement as indicated by a 180" bend around a diameter approximately equal to the diameter of the wire (0.125 in.) : General solution of the surface had taken place, indicated by a change in diameter of the wire, and the resulting structure of the surface is shown in Fig. 1. It is apparent that angles have developed at grain boundaries and examination at a high magnification indicates that the angle is in excess of 120°, completely excluding penetration of mercury. In view of the measurements made by Smith n the

Jan 1, 1952

By Fredrick C. Pearson, Albert Brown, Emil R. Ermert

EARLY in 1946 the Shen-Penn Production CO., a subsidiary of the Philadelphia and Reading Coal and Iron Co., was organized to operate the Shenan-doah Stripping, one of the largest open-pit anthracite mines in the world. This mine is located in the Western Middle Field near Shenandoah, Schuylkill County, Pa., and involves the removal of approximately 60,400,000 cu yd of overburden to permit the recovery of 9,200,000 gross tons of rough cleaned coal from an area 8000 ft long and 2700 ft wide, with a maximum depth of 400 ft. Operations began in August 1946, and within the next few years 31,811,-000 cu yd of overburden and 2,994,000 gross tons of rough cleaned coal had been removed. During the early days of open-pit mining operations, steam was the primary motive power. However, as the size of the equipment increased, so did the problems of fuel, water, labor, and maintenance. Because of its efficiency and convenience, electric power rapidly replaced steam. Later the modern diesel engine was developed, and today practically all stripping equipment is either electric, diesel-electric, or diesel-powered. In locations where electric power is not readily available or where operations are scattered over a large area, diesel power is favored. At Shenandoah operations are highly concentrated in a relatively small area, and this fact, coupled with the availability of ample electric energy, indicated the selection of electric-powered equipment. Since this was a new company, it was necessary to purchase every item, from the smallest tools to stripping shovels. Although the organization period was hampered by the postwar scarcity of materials, an excellent opportunity was afforded to set up a thoroughly modern plant. This paper will describe the power facilities, with particular emphasis on the more unusual features. The substation is a major item in a power distribution system, and a simple arrangement utilizing high quality equipment is an excellent investment. With this in mind, a 3000-kva unit type substation was selected and installed. This unit is made up of three sections, which are assembled simply by bolting together the various sections and connecting the power leads, see Fig. 1. The transformer proper, rated at 3000 kva, is oil-filled and nitrogen-sealed, having a 66,000-v Delta primary and 4160-v Wye secondary with a grounded neutral and 46-ohm ground resistor. The function of this resistor will be described later. Electrical energy is purchased from a public utility and is delivered to the main substation at 66,000 v over one or the other of two incoming lines. These lines terminate on a high voltage switching section on which high voltage fuses and Thyrite lightning arresters are installed. The outgoing feeder switching unit is made up of the following: 1—An auxiliary compartment consisting of potential and current transformers with necessary terminal blocks for connecting to power company metering equipment. 2—Another auxiliary compartment containing a control transformer, battery, and battery charger for the furnishing of power for tripping and reclosing of power circuit breakers. The assemblies in these auxiliary compartments are mounted on carriages that can be rolled in and out with ease. Potential transformers are readily accessible for inspection and fuse replacement. When withdrawn from the operating position they are completely de-energized and grounded. 3—Two feeder breaker compartments, each with a 1200-amp, 5000-v, solenoid-operated magneblast air circuit breaker, voltmeter, ammeter, overcurrent relays, and a ground relay. These power circuit breakers can be quickly and safely withdrawn or repaired. A built-in isolating mechanism is used for lowering the breaker to a transfer truck for repair or inspection or as a means of disconnecting the breaker from the source of power as an added precaution when working on power lines. Thus inspec-

Jan 1, 1953

By Albert Brown, Emil R. Ermert, Fredrick C. Pearson

EARLY in 1946 the Shen-Penn Production CO., a subsidiary of the Philadelphia and Reading Coal and Iron Co., was organized to operate the Shenan-doah Stripping, one of the largest open-pit anthracite mines in the world. This mine is located in the Western Middle Field near Shenandoah, Schuylkill County, Pa., and involves the removal of approximately 60,400,000 cu yd of overburden to permit the recovery of 9,200,000 gross tons of rough cleaned coal from an area 8000 ft long and 2700 ft wide, with a maximum depth of 400 ft. Operations began in August 1946, and within the next few years 31,811,-000 cu yd of overburden and 2,994,000 gross tons of rough cleaned coal had been removed. During the early days of open-pit mining operations, steam was the primary motive power. However, as the size of the equipment increased, so did the problems of fuel, water, labor, and maintenance. Because of its efficiency and convenience, electric power rapidly replaced steam. Later the modern diesel engine was developed, and today practically all stripping equipment is either electric, diesel-electric, or diesel-powered. In locations where electric power is not readily available or where operations are scattered over a large area, diesel power is favored. At Shenandoah operations are highly concentrated in a relatively small area, and this fact, coupled with the availability of ample electric energy, indicated the selection of electric-powered equipment. Since this was a new company, it was necessary to purchase every item, from the smallest tools to stripping shovels. Although the organization period was hampered by the postwar scarcity of materials, an excellent opportunity was afforded to set up a thoroughly modern plant. This paper will describe the power facilities, with particular emphasis on the more unusual features. The substation is a major item in a power distribution system, and a simple arrangement utilizing high quality equipment is an excellent investment. With this in mind, a 3000-kva unit type substation was selected and installed. This unit is made up of three sections, which are assembled simply by bolting together the various sections and connecting the power leads, see Fig. 1. The transformer proper, rated at 3000 kva, is oil-filled and nitrogen-sealed, having a 66,000-v Delta primary and 4160-v Wye secondary with a grounded neutral and 46-ohm ground resistor. The function of this resistor will be described later. Electrical energy is purchased from a public utility and is delivered to the main substation at 66,000 v over one or the other of two incoming lines. These lines terminate on a high voltage switching section on which high voltage fuses and Thyrite lightning arresters are installed. The outgoing feeder switching unit is made up of the following: 1—An auxiliary compartment consisting of potential and current transformers with necessary terminal blocks for connecting to power company metering equipment. 2—Another auxiliary compartment containing a control transformer, battery, and battery charger for the furnishing of power for tripping and reclosing of power circuit breakers. The assemblies in these auxiliary compartments are mounted on carriages that can be rolled in and out with ease. Potential transformers are readily accessible for inspection and fuse replacement. When withdrawn from the operating position they are completely de-energized and grounded. 3—Two feeder breaker compartments, each with a 1200-amp, 5000-v, solenoid-operated magneblast air circuit breaker, voltmeter, ammeter, overcurrent relays, and a ground relay. These power circuit breakers can be quickly and safely withdrawn or repaired. A built-in isolating mechanism is used for lowering the breaker to a transfer truck for repair or inspection or as a means of disconnecting the breaker from the source of power as an added precaution when working on power lines. Thus inspec-

Jan 1, 1953

By Klaus Detert, Joseph Ziebs

The progress of primary recrystallization has been studied in alloys of nickel with 0.01, 0.1, and 1.0 at. pct Ta. It can be shown that in one particular alloy containing 0.1 Ta and 0.02 02 a large grain structure (0.5 mm diameter) could he developed by primary recrystallization in thin sheetswhich were cold-worked to 97.5 pct reduction in thickness. Electron transmission microscopy demmstrates that this phenonmena was linked to the occurrence of fine precipitates at subboundaries. A discussion of how fine precipitates can exert under certain circumstances a rather distinct retarding effect on nucleation rate is included. IN a former investigation, the recrystallization and texture formation in a series of binary nickel alloys including pure nickel were studied.' Nickel alloys with some tantalum added displayed peculiarities during primary recrystallization which have been studied in more detail by the present investigation. The alloys were prepared by powder-metallurgy techniques similar to the previous investigations.' Pure metal powders were thoroughly mixed to obtain alloys with 0.01, 0.1, and 1.0 at. pct Ta, referred to as alloy A, B, or C, respectively. The metallic impurities were kept below 0.005 at. pct except iron and cobalt with slightly higher values, by using highly purified nickel and tantalum powders.' The mixtures were pressed into bars and sintered at a temperature of 1300°C. The sintering atmosphere was either high vacuum, dry hydrogen, or wet hydrogen. The sintered rectangular bars were rolled 60 pct to a thickness of 3 mm, then annealed under dry hydrogen at 1000°C and again cold-rolled to about 97.5 pct reduction of thickness. After this heavy deformation, the sheet thickness was 0.07 5 mm. In order to study primary recrystallization which occurred in measurable times between 400" and 600°C, samples of the rolled sheets were annealed in a salt bath at various temperatures for various times in order to study the progress of recrystallization by hardness measurements and microscopic examination. For optical observation, the samples were polished and etched in a solution of 65 cu cm HNO3, and 18 cu cm acetic acid in 20 to 100 cu cm H2O, depending of the particular sample and annealing treatment. For transmission electron microscopic observations, thin foils were prepared by the combined window and Bollmann techniques.273 The electrolyte consisted of 23 pct perchloric acid and 77 pct acetic acid held at 15°C. Polishing was performed at a current density of 0.4 amp per sq cm. The examination of the thin films was accomplished by a Siemens Elmiscope-type electron microscope operating with an accelerating voltage of 100 kv. RESULTS The progress of primary recrystallization was very accurately followed in these alloys in order to establish the time when recrystallization started and ended during an isothermal anneal. Table I gives some of the data obtained by these studies. The time for beginning is a mean value, where somewhat less than 5 pct of the matrix has recrys-tallized. The end of recrystallization is the time when at least 95 pct of the matrix has recrystal-lized. Listed are three different B alloys, with 0.1 Ta distinguished by different sintering atmospheres which influence the rate of primary recrystallization. By the above method, the state of the grain structure when primary recrystallization was completed, but grain coarsening had not started, could be clearly defined. The average grain diameter of the recrystallized matrix is also given in Table I. The grain size did not depend on the temperature of the isothermal anneal. With one exception the grain size of the primary recrystallized matrix was found to be rather fine

Jan 1, 1965

By H. M. Myers, J. E. Funk

The two-phase flow of a drilling fluid in a diamond drill bit is investigated by deriving and solving the steady-state continuity and momentum equations. Parameters in this analysis are rotational speed of the bit, local rate of addition of drill solids, non-Newtonian nature of the drilling fluid and drilling rate. The describing equations relate the flow and pressure drop to the design characteristics of the bit. The equations are integrated numerically yielding the flow-rate pressure drop characteristics for the bit, flow distribution in the bit and velocity and static pressure as a function of position in each flow channel of the bit. Results are presented for a particular bit design. This paper shows that under normal drilling conditions certain parameters may be neglected in the hydraulic calculations for a diamond drill bit. Therefore, shortened forms of the equations presented can be used to fit the proper diamond bit to the drilling fluid hydraulic system. INTRODUCTION Like the conventional bit, the diamond drill bit is a key component in the circulating system of a well. The conventional rock bit is fitted to the circulating system by selecting the flow nozzles which provide the system with the proper flow-rate pressure drop relationship. Fluid behavior across the face of the diamond bit must be better understood so the bit can also be fitted to the circulating system. Data from field-operated diamond bits are presently available and do yield some information regarding the flow-rate pressure drop relationship. However, the number of factors affecting this relationship is quite large. Drilling parameters and conditions such as bit penetration rate, bit weight and rotational speed are important. Moreover, flow channels in the bit are part of the cutting head, and introducing drilled solids into the flow channels reduces the flow area, causes additional drag and influences the pressure drop. Many tests would have to be performed if field tests were required to study all parameters which might influence fluid behavior in the bit. A logical approach to this problem is to investigate analytically the fluid behavior in the bit, including parameters that are present under normal drilling operations. This approach, which seems to offer the best method for future studies, is presented in this paper. Following is the general procedure. 1. Equations describing the most general flow channel are written and include the addition of drilling solids (which is related to penetration rate), rotation of the bit, design parameters of the bit and non-Newtonian properties of the fluid. 2. A typical bit is designed and the bit parameters are substituted into the equations. 3. Parameters are assumed (penetration rate, rotational speed of the bit, mud weight, etc.). 4. The describing equations are numerically integrated to obtain the pressure drop across the face of the bit as a function of flow rate. 5. The pressure-drop contribution of each component (friction, acceleration, etc.) is obtained from the numerical solution, and those factors which largely influence the total pressure drop are identified. Classifying the drilling fluid is important in establishing the friction factor relationship since the frictional pressure losses are the main contributors to the total pressure drop. High-density drilling fluid is commonly assumed to be a Bingham plastic. 1-3 Govier4 presented a method for selecting a friction factor in the laminar flow region from the rheology properties of the mud and a modified Reynolds number. Metzner5 presented a criterion for selecting the friction factor in the turbulent flow region. The bit design used in conjunction with the solution of the equations has divergent flow channels. Stall, which is the backflow of fluid in the channels due to eddies that occur on the outer perimeter of the flow stream, can occur in these divergent channels (or diffusers). When stall does occur the static

By R. L. Slobod, B. H. Caudle

In the past, the main emphasis in attacking the problem of the recovery of oil has been on the determination of fluid flow characteristics and residual oil saturation in the part of the reservoir which is contacted by a displacing phase. It is recognized that the determination of the fraction of the reservoir contacted in secondary recovery operations is also of great importance in predicting the ultimate recovery to be expected. A method is described in which radiographic techniques are used to determine these areal sweepout factors for any type of well spacing. Data showing the relationship of viscosity ratio and relative permeability to areal sweepout efficiency are presented for the five spot and the direct line drive well spacing patterns. Further applications of the method to studies of the displacement efficiency in the area contacted are mentioned. INTRODUCTION The problem of increasing oil recovery from typical reservoirs is generally attacked by trying to devise some means for reducing the residual oil in that part of the reservoir which is contacted. The fraction of the reservoir contacted in the displacement process. however, is also of major importance but has been given less consideration in the past. one reason for this neglect is the meager information available on the portion of the reservoir contacted and the magnitude of changes in this factor as operating conditions are varied. Until recently, the only means available for determining this quantity has involved the use of the electrolytic or poten-tiometric modelsl in which one is limited to working with a mobility ratio of unity. while means for calculating the volume of a reservoir contacted do exist. the calculations are very tedious for systems with mobility ratio different from unity. Recently several such determinations have been reported,213 and an experimental method involving a step-wise use of potentiometric models has been described.' A new method which has been developed for studying the areal distribution of phases in a porous medium is discussed below. This tool can be' used to determine the effect of operating variables such as viscosity, relative permeability, presence of gas, mobility ratio, rate and stratification on the fraction of the reservoir contacted by the displacing medium. This development based on the use of X-rays gives a pictorial representation of the location of the phases. Further, the amount of each phase at each point of interest can be determined quantitatively. X-RAY METHOD FOR AREAL STUDIES In this method the location of various phases distributed Over a test area are observed simultaneously by obtaining a photographic record of the transmission of X-rays through the test plate. This procedure is an adaptation of the radiographic techniques which have 10% been used in the medical and industrial fields. Just as the medical profession uses a chest plate to show spatial interrelationships, the location of an X-ray absorbing fluid and a nonabsorbing fluid in a porous medium can be determined by obtaining a photographic record of transmitted X-rays. Moreover, as changes occur in such a System, these changes can be observed by taking successive photographs. Previously, X-ray methods have been used in fluid flow studies to determine quantitatively the saturation at a given point in the system. The intensity of the transmitted X-rays was measured by an ionization chamber, so that a point by point survey of saturation could be made. In comparison with this earlier quantitative procedure, the photographic procedure which simultaneously observes the saturation at many

Jan 1, 1952

By L. W. Strock

Metallurgists handling lead and zinc ores have long been familiar with the spectrograph as a routine analytical tool, as its earliest regular use by American industry was in controlling impurities of zinc metal and its alloys. This application has been extended to other metal industries on an extensive scale, accompanied by refinements of procedure to such an extent that a large part of some metals produced, especially magnesium, aluminum and steel, are checked only by spectrographic methods for conformity to specifications. The spectrographic analysis of ores and other poorly conducting powdered solids presents difficulties not encountered with metallic samples, which can be arced or sparked as self-electrodes or can be readily dissolved, so that their solutions can be analyzed, as in the American Society for Testing Materials standard methods for zinc and zinc-base alloys. Spectra of powdered solid samples can be prepared by using carbon electrodes provided with a crater of suitable size, into which the sample is packed. Ore samples and many other powdered solid materials usually analyzed in the chemical and metallurgical laboratory are more complex and varied in composition than samples of metals produced in large tonnages. 'lore preliminary work therefore is usually necessary in devising a spectrographic method for analyzing a given type of powdered sample than for analyzing metal samples, in order to use that method with confidence. The howledge that the standard samples and unknowns produce equivalent line intensities at equal concentration is the fundamental basis of any quantitative spectrographic method. On the other hand, powdered samples have an advantage over metals in that any desired number of chemical compounds can be mixed dry in any desired ratio to produce a synthetic standard. Expcrience has shown, however, that spectra of elements in finely divided compounds introduced into a standard in preparing a synthetic ore sample do not always behave in the electric arc in a manner that produces spectral line intensities similar to those observed at the same concentration in a natural mineral. In ore samples minor elements frequently are contained in the crystal lattice of larger particles of an abundant ore mineral. Their dissociation and volatilization from the sample into the arc then proceeds differently in the standard and sample, with resultant differences in intensity. During the past 10 years I have devised and applied several quantitative spectrographic methods for determining various elements in rocks and minerals in connection with studies on their geochemistry,' but the incentive has gradually changed since the summer 0. 1940 to the specific practical problem of searching for domestic sources of rare and stratcgic metals. The

Jan 1, 1949

By K. E. Nelson

The effect of thorium and zinc variations on the strength and 100-hr creep characteristics of Mg-Th-Zn-Zr alloys was investigated. Optimum resistance to creep at 650° and 700°F are attainable within a certain range of thorium and zinc contents. This range does not conform to that which develops maximum tensile properties. RAPID advances have been made during the last few years in the development of magnesium alloys for elevated temperature applications demanding high resistance to creep. The beneficial effect of rare-earth metals on the creep resistance of magnesium alloys has been emphasized by a number of publications1-13 and such alloys are now in commercial production. The use of thorium as an alloying ingredient in magnesium was mentioned by McDonald and also in two Alien Property Custodian patent applications.10,17 The initial observation of Sauerwald that thorium contributes still higher creep resistance to magnesium than is attainable with rare-earth metals has recently been substantiated.'" " In fact, it has been demonstrated that the useful temperature range of magnesium alloys is appreciably extended by the use of thorium. In all cases, it was observed that zirconium must be included in the alloys in order to render them fine grained and more readily castable. Several recent publications:2-27 indicate that a still further improvement in creep resistance and a further extension of the useful temperature range can be realized by the addition of zinc to alloys. The primary purpose of this paper is to present the results of a comprehensive study of the effect of zinc on the strength and creep characteristics of Mg-Th-Zr alloys. Compositions covering the range of thorium content from M to 6 pct and zinc content from 0 to 5 pct have been investigated. The creep characteristics at 650" and 700°F reported in this paper are based on results of tests of 100-hr duration. It is appreciated that creep tests of 100-hr duration might not yield adequate data for design purposes for parts with much longer expected life. However, for the purposes of the present discussion, it is felt that the combination of stresses and temperatures used in the 100-hr creep tests have yielded a clear representation of the compositional variation of creep resistance at the temperatures investigated. Creep tests of 1000-hr duration are now in progress on a few of the most promising alloys. Preparation and Testing of Alloys The alloys studied in this intensive investigation were prepared in 25 lb capacity mild steel crucibles. The thorium, zinc, and zirconium were alloyed and poured as described in earlier publications The thorium was introduced into the melt in the form of a Mg-Th hardener," the zinc added in the metallic form, and the zirconium alloyed in the form of the commercial hardener containing magnesium and 30 to 50 pct Zr.10, 26 Fluxing practices for melting and refining were the same as for magnesium-rare-earth metal-zirconium alloys. The melts were sampled for analytical determinations and poured into separately cast 1/2 in. diameter standard tensile bars. The test bars were given a precipitation treatment of 16 hr at 600°F in laboratory furnaces. It has been shown by other tests that a high temperature solution treatment followed by an aging treatment is unnecessary for the development of optimum properties in Mg-Th-Zn-Zr alloys. The selection of 600°F as the aging temperature was based on an attempt to achieve metallurgical stability without coalescence of the undissolved phases and the attendant loss in strength. The thorium, zinc, and zirconium contents of each melt were determined chemically. The zirconium contents are reported in two parts, "soluble" and "insoluble," referring, respectively, to the portions present in the alloy which are soluble and insoluble in dilute HCl acid. Distinction is being made between these two components of the zirconium content in the alloys because it has been found that only that portion of the zirconium content which is soluble in dilute inorganic acids affects the structure and properties of the alloys. The usual impurities consisting of copper, iron, manganese, and nickel were determined spectroscopically. The analysis for each melt is listed in Table I. A description of the methods of tension and creep testing has been detailed in earlier papers. The tests were performed with the cast skin re-

Jan 1, 1954

By Joe B. Alford

Jan 1, 1971

By A. U. Seybolt

The solubility of oxygen in a iron has been determined in the range between 700° and 900°C. The solubility is a function of temperature and varies from about 0.008 pct oxygen at 700°C to atureandabout 0.03 pct at 900°C. The heat of solution is approximately +15,500 cal per mol. AS pointed out in a recent paper by Kitchener et al.,1 there has been a lack of agreement among many investigators even as to the order of magnitude of the solid solubility of oxygen in the various forms of iron. This lack of agreement is attributable in large part to the difficulties in the determination of a small oxygen solubility; but because the problem has remained so long unsettled, it also indicates a lack of interest which is rather surprising when the demonstrated importance of small amounts of soluble nonmetallic impurities in iron is considered. The work of Kitchener et al. apparently leaves the solubility of oxygen in iron in a satisfactory state, but no attempt was made to investigate the solubility in a iron. The solubility of oxygen in a iron is actually of greater interest, since it is in this form that iron and mild steels are employed ordinarily. That the effect of oxygen in iron is of more than theoretical interest has been well established by Fast,' and more recently by Rees and Hopkins," who demonstrated that oxygen in the range between 0.0008 and 0.27 wt pct has a pronounced effect upon the mechanical properties. Previous Work and Methods Used To report in detail the literature relating to the solubility of oxygen in a iron would require an inordinate amount of space. For those interested in reviewing this work, a bibliography4-13 of the more significant papers is appended. In general, two methods of studying this problem have been used. One is the gas-metal equilibrium method where the H2O-H2-Fe or the CO2-CO-Fe equilibria have been used. The other is the more direct approach of the oxidation of thin strips of pure iron by packing in mill scale or by air or gaseous oxygen at some desired temperature. In this method oxygen is allowed to oxidize the surface and then to diffuse inward until saturation is obtained. In the gas-metal equilibrium method the oxygen dissolved in solid iron at a given temperature is proportional to the ratio of the water vapor-hydrogen pressures or the CO2-CO pressures over the sample for small ratio values. If the ratio becomes higher than a critical value, then an oxide phase makes its appearance (the solution becomes supersaturated). In principle, it is possible to use a series of H2O/H2 or CO2/CO ratios and to find by analysis the corresponding amounts of oxygen in solid solution at constant temperature. At the point where a very small increase in the gas ratio (increase in oxidizing powder) produces a large increase in oxygen content, the solid solubility limit is reached. Alternately, if the critical ratio is known, it is possible to use the procedure of Kitchener et al.1 and to use a ratio which is near but below the critical one. The solubility corresponding to this lower ratio will not be the saturation solubility at the temperature employed, but the saturation solubility can be calculated by multiplying the measured solubility by the critical ratio over the ratio used. However, in the case where oxygen gas is the oxidizing medium, the saturation solubility is not a function of pressure, providing the pressure exceeds the dissociation pressure of FeO in equilibrium with iron. This is 1.2x10-16 mm at 800 °C, according to Dushman." As pointed out by Darken17 in discussing the FeO phase diagram, most of the possible errors tend to yield high values of oxygen solubility. For example, one circumstance which evidently caused the reporting of many high values was the use of finely divided or powdered iron in the gas equilibrium method. Because of the large surface area of such a sample, and the likelihood of some surface contamination if only by exposure to air, the results tended to be high. The direct oxidation method which was used in this work has the advantage in that it is simple and direct, but it suffers from one disadvantage: equilibrium can only be approached from the low oxygen side. The important factors to be kept under control are the following: 1—use of high purity iron to avoid internal oxidation (oxidation of readily

Jan 1, 1955