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By C. E. Gerity

Near Bunker, Mo., in the New Lead Belt, Ozark Lead Co. operates a 6000 tpd lead-zinc mine. Mined ore is transported to the surface in three mechanized and automated steps. The mine is developed by a rectangular-shaped sublevel haulage drift 15 ft high by 28 ft wide, driven on strike below the ore zones. Ore transfer raises were driven from the sublevel at locations near the shaft, 1500 ft distant, and on subsequent 1000 ft centers. These 6 x 5-ft raises are slashed to obtain maximum storage below dump points. These dump points are located on each mining level and are of steel and concrete construction. Spiral 12 X 24-ft inclines driven from the sub-level provide access to these levels for equipment and mine utilities.

Jan 11, 1972

NEW MEMBERS. The following list comprises the names of those persons who became members during the month of April, 1913: BARNES, BLAKESLEE, Min. Engr Burrwood, La. BOYLSTON, HERBERT M., Cons. Met. Engr 26 Abbot Bldg., Cambridge, Mass. BUEHLER, HENRY A., State Geol Rolla, Mo. CASTO, WILLIAM H., JR., Mining Bonanza, Idaho. COHEN, MAXWELL B., Mill Operator Miami, Ariz. EVANS, MILTON R., Shop Draftsman 92 Division St., Kingston, Pa. HARPER, WALTER S., Asst. Treas 42 Broadway, New York, N. Y. JAMES. JAMES S.. Min. Engr Telluride. Colo.

Jan 5, 1913

By A. A. Toprac

The recurring problem of joint failures in ofjshore structures, coupled with the catastrophic collapse of many of these structures during Hurricane Hilda (1964), hay forced reconsideration of techniques presently used to design sfuctural joints. This paper presents results of studies conducted for this reconsiderat?on. Only right-angled T-joint,s and cross-joints where a brace was welded directly to the undisturbed surface of another tube (the chord) were considered. Tests were performed on seven T-joint and 12 oross-joint specimens of two types which were designed to give a broad but practicd range to the geometrical parameters believed to be of significance to the problem. These paramefers were expressed in dimensionless form. Each specimen of the T-joints was instrumented wirh both electrical resistance strain gauges and deflection dials. After the elastic tests were concluded, each specimen was loaded to failure by applying an axial tensile load to the brace. The cross-joints were tested to failure without measuring elastic strains. Theoretical solutions for the stresses in a cylindrical shell due to a distributed radial load over portions of its surface were numerically evaluated. The experimental results are compared with calculated values obtained using Bijlaard's and Dundrova's theoretical analyses. INTRODUCTION During the last IS years a large number of structures were built for the offshore exploration of oil both in the Gulf of Mexico and dong the Pacific Coast.' In all these offshore platforms, the tubular shape was extensively used since this shape provides outstanding strength in proportion to its weight, compared to other shapes, when all directions (x, y, z) are considered. Past use of tubes was hampered because of connection details. But with the advent of welding, it became no more difficult to connect tubular shapes than to connect rolled shapes. However, some problems have been encountered in actual construction in making welded joints that are adequate and yet economical. In trying to connect tubular members, designers have used a varied number of methods; consequently, some of the joints seem to be very expensive in detail while others appear weak and underdesigned. An inspection of offshore platforms owned by different companies indicates the use of a large number of joint arrangements and configurations. Such a variety of joint types in current service indicates the wide divergence of opinion that exists on how such joints should be designed, stiffened and strengthened. Fig. 1 shows the various types of connections that have been used in practice. This paper reviews the present state of art on tubular joints of the T and cross type, and presents results of experimental and analytical work being conducted at The U. of Texas and elsewhere. It is believed that such information will be of great help to designers for the formulation of proper design procedures for joints. CONNECTION PROBLEM Recently, cracks have been observed in several welded connections in offshore structures. Fortunately, until about 1 1/2 years ago, these failures have been localized and have not resulted in the collapse of structures or loss of life. For example, one platform installed in Louisiana during 1955 in about 60 ft of water had been subjected to four tropical hurricanes. This platform successfully withstood these forces and was later removed because operations at this location had been completed. Many joint failures were observed during the removal.

By I. H. Van Horn

IN the development of the incandescent electric lamp one aim of the investigators has been to establish the fundamentals of lamp design, so that the performance of any new lamp may be accurately predicted. The study of the temperature relations in lamps has done much toward establishing these fundamentals. Fig. 1 gives the typical FIG. 1.-ARRANGEMENT OF PARTS IN MAZDA B AND MAZDA C LAMPS. arrangement of the lamp parts in both the Mazda B (vacuum) and Mazda C (gas-filled) lamps. The outstanding differences are the filament form and the bulb shape. The ideal lamp is one in which the filament operates at a uniform temperature throughout its length. In practice, it is necessary to disturb this uniformity by the introduction of supports and terminals. The amount of cooling at the supports and terminals affects the over-all efficiency of the filament as a light producer. The life of the lamp is ordinarily proportional to the maximum and not to the average filament temperature. The Holborn-Kurlbaum type of optical pyrometer has

Jan 9, 1919

By Hubert E. Risser

Although, for about a half-century the percentage of coal production provided through strip mining has steadily increased, recent trends indicate that a peak in percentage (not tonnage) is being approached, principally because of the decline in economic adrantage enjoyed by strip mining compared to underground mining. Factors contributing to the decline are the decrease in the margin of advantage in manpower requirements held by strip mining; an increase in depth and volume of overhurden, requiring larger equipment and capital investment; increasing lack of reserve land blocks large enough to support extensive strip mining operations; more stringent legislation for reclamation of stripped land, with increasingly higher costs of such reclamation; and the increasing costs of land surface rights for strip mining. Already there has been a peak and suhsequent decline in relative importance of strip mining insome major Eastern coal-producing states; trends indicate that a peaking for the whole Eastern region is on the way. Although, in the West the pressures are less severe, they will, nevertheless, reduce to an increasing degree the advantage that strip mining has possessed. For about half a century, strip mining has been providing an ever-increasing share of total coal output in the United States, and in the past 30 years this share has grown from less than 8% to 33.7%, as shown in Fig. 1. Fig. 1 also shows that strip mining is gaining faster in the West than in the East. From 1961 to 1967, output from strip mines grew 63.6 million toqs, or 52.2%. Certainly there is nothing in such performance to indicate any peaking in the level of strip coal-mine production. But the fact is that, if not in actual tonnage, strip mining appears to be reaching its peak in relative importance, at least east of the Mississippi River. And the area east of the Mississippi provides 90.0% of the strip and 94.9% of the total coal production of the nation. The relative economic advantage which spurred strip mining's growth for a third of a century has been diminishing markedly in recent years. Fig. 2 shows evidence of this decline. From 1947 until 1961, strip tonnage output remained fairly stable, at the end of the period showing only a 12.5% decline. Meanwhile, underground coal production fell 44.5%. As the result of the greater decline in underground coal, strip coal in 1961 furnished 30.3% of the total coal produced, compared to only 22.1% in 1947. In the succeeding six years, however, underground coal output increased by 76.5 million tons, while strip output increased only 63.6 million. Further evidence of the strengthening of underground coal's relative position — or if you will, the weakening of strip coal's position — can be found in the major coal stripping states: Illinois, West Kentucky, Ohio, Pennsylvania, Indiana, and West Virginia, which in 1966 provided 80% of the nation's strip coal. Fig. 3 shows what has been happening in three states lying in the Appalachian field and in the total output of these three states. In the figure, the column height represents the output of coal mined by all methods. The number at the base of each column is the percentage mined by stripping. In Pennsylvania and Ohio, the percentage of strip mining increased between 1956 and 1961, but by 1966 had declined from the 1961 level. In West Virginia, a decline occurred during the first five years but was followed by a rise. In January 1968, Coal Age published an article listing the locations and capacities of new large mines in the planning and construction stages that would be operating by 1968-1971. The plans for West Virginia and Pennsylvania indicated all the new mines would be underground. Except to the extent they might replace existing under-

Jan 1, 1970

WE the undersigned, being all persons of full age and citizens of the United States, and a majority residents of the State of New York, desiring to form a corporation pursuant to the provisions of the Membership Corporations Law for the purpose of incorporating, as provided in Section 6 of Article I. of said law, the existing unincorporated association known as American Institute of Mining Engineers, do hereby make, acknowledge and 6le this Certificate for that purpose, and do Certify as follows: I. That the American Institute of Mining Engineers is an unincorporated association organized and existing with the object of promoting the arts and sciences connected with the economic production of the useful minerals and metals and the welfare of those employed in these industries by means of meetings for social intercourse and the reading and discussion of professional papers, and to circulate by means of publications among its members and associates the information thus obtained. 11. That the persons duly appointed or designated to manage the affairs of said association are designated by the rules thereof Members of its Council; that the under- signed are all members of said Council as the same was constituted on the 29th day of December, 1904.

Jan 1, 1942

By Samuel M. Shallcross

The principal function of the Bell limestone mine, of the American Lime & Stone Co., at Bellefonte, Pa., is to supply raw material to the company's modern rotary kiln plant at Bellefonte. Because mining costs are naturally higher than many open quarries with more favorably lying strata, it cannot compete, except to a limited extent, in the markets for road material and fluxing stone. In the broad sense, mining does have an advantage over the average open quarry when the end products are designed for and controlled by process and final inspection for the exacting demands of the industrial and chemical markets. Approximately 80 per cent of the output of the Bell mine is sold in the industrial and chemical markets. And as higher standards are set each year for construction and agricultural work, the remainder of the mine output goes to these two classes. Mining produces more fines than open quarries, and these fines are difficult to market; on the other hand, owing to absence of contaminating clay and impure rock, a greater percentage of the output of a mine is marketable. Because of the great production of fines, rotary kilns for lime burning are necessary when the stone must be mined. In 1921 the American Lime & Stone Co. operated 19 quarries and 72 lime shaft kilns spread over an area of 40 miles between Bellefonte and Altoona, Pa. At that time the lime-production capacity of the 72 kilns amounted to approximately 700 tons per day. After 15 years the company operates three rotary kilns and six shaft kilns, with a combined output of approximately 500 tons of lime per day, concentrated in its single Bellefonte plant (Fig. 1). Its only stone-producing plants are now the limestone mine at Bellefonte and a construction-stone open quarry at Union Furnace, near Tyrone, Pennsylvania. Since the middle of 1922, the management of the American Lime R. Stone Co. has been under the direction and ownership of the Warner Company, of Philadelphia. To the Warner Company's experience, gained through successful operation of lime plants in eastern Pennsylvania for many years, and the successful solution of material-handling problems

Jan 1, 1937

By R. S. Olson

An analysis of the wash thickener-two stage filtration circuit was undertaken to quantitatively describe the circuit performance. On the basis of information developed in this study, the factors of major importance in reducing soluble loss are: filter wash, wash zone size, filter capacity, filter feed pulp dilution, blow back, filter cloth, water to repulp, water to primary wash, water to feed pulp and thickener underflow. The wash thickener-two stage filtration circuit has been used in many hydrometallurgical plants to separate solute from leach solids. As many as 20 to 30 plant operating variables determine the performance of the wash circuit and hence the soluble loss. The complex relationships of these variables, with each other and with soluble loss, make the evaluation of their effect on mill performance difficult. An analysis of this type of circuit was undertaken to quantitatively describe the circuit performance. THE CIRCUIT The diagram in Fig. 1 represents the operation of the wash thickener-two stage filter circuit. This circuit provides for counter current, cross current or for a combination of countercurrent-cross current operation. The feed pulp stream, J, is intended to include any solution or water, other than primary and secondary filtrate, which enters the wash thickener. Except for filter washing, the operations in the circuit are simple mixing operations which can be described by material balance. The concentration of solute in the filtrate and cake solution after application of wash liquor cannot be described by simple material balance. The displacement washing equation, derived by Rhodes1 and extended by Choundhury and Dahlstrom,2 appears to best describe this case of miscible displacement in porous media. In a useful form, it is

Jan 1, 1963

By J. C. Osmond

One of the least known geologic regions in the U.S. is the area now called the Basin and Range Province. It is paradoxical that so little geologic information has been compiled for a province that has yielded billions of dollars worth of metals and is currently producing a small amount of oil. Regional stratigraphic studies of parts of this large territory have been published only recently, and regional structural interpretations of even parts of the province are rare in the literature. This, of course, does great injustice to the earliest workers: Gilbert,1,2 Russell,9 Le Conte,4 Spurr,5 Louderback,8 and others, who recognized the area as distinctively different from the rest of North America. Their simplified picture of the structure of the ranges and the province as a whole is the groundwork for present understanding and has never been superseded. Since their early works, however, geologic data from the province has consisted largely of detailed reports of widely scattered mining districts. Each district has been described as though it were an island in an unknown sea. More recent syntheses of the geologic history have been prepared by Lindgren,7 Ransome,8 Nolan,9 Bill-ingsley and Locke,30 Longwell,11 and Eardley.12 The tentative nature of the present interpretation, which treats the structural evolution of the Utah and Nevada area, must be considered in light of the scarcity of geologic information from the province as a whole. The methods of interpreting tectonic history from stratigraphy are based largely on the concepts of Marshal Kay.23 Discussions with Hoover Mackin also have aided in organizing the Tertiary history. TECTONIC AND STRATIGRAPHIC SETTING Two approaches to the subject of the tectonic history are possible; they correspond to looking through opposite ends of a pair of binoculars. One involves an expansion and extension of detailed work and the other an interpretation based on regional geological relationships. It is believed that the second approach, used here, is not in disagreement with detailed observations. The area of the Basin and Range Province is shown in Fig. 1. Within this area middle Tertiary to Recent high-angle faulting has controlled development of mountain ranges and valleys. Only the part of the area north of Las Vegas, Nev., and south of the Snake River in Idaho will be considered in this article. Within this part of the province most of the ranges trend north-south. Approximately two thirds of the area is covered by late Tertiary to Recent valley fill and lake deposits.

Jan 1, 1961

Jan 1, 1964

By S. R. B. Cooke, H. S. Choi, I. Iwasaki

In a previous article1 the function of various fatty acids as collectors for iron ores was reported for the two alternate processes; (a) the flotation of iron-oxide minerals, and (b) the flotation of calcium-activated siliceous gangue. The selectivity of separation was shown to be markedly dependent not only on such common parameters as pH and the levels of collector and activator addition, but also on the structure and the degree of unsaturation of the acids used. The iron ores employed consisted primarily of hematite, goethite, and quartz and as the effectiveness of any one acid on the separation depends on the individual response of each constituent mineral toward the given collector, a comprehensive knowledge of the floatability of the individual minerals is desirable for the interpretation of the results obtained. The mechanism by which a collector ion becomes attached to a mineral-solution interface has been a subject of much investigation. From quantitative studies of the properties of the electrical double layer on mercury2 and on several solid surfaces3,4,5,6 , the adsorption of some flotation reagents at the mineral-solution interface is beginning to be understood. For example, it has been demonstrated from streaming potential measurements made on quartz5 and corundum6 that the interfacial electrical condition is strongly dependent on the pH. Thisoccurs presumably through the following electrolytic reaction at the interface, and in the presence of a flotation collector, controls the floatability of the solids. A concurrent investigation of the flotation behavior of goethite and of quartz carried out at the School of Mines and Metallurgy7 has shown a positive correlation with the electrokinetic behavior of these minerals. The purpose of the present investigation was to collect general information on the floatability of hematite, and to investigate the function of various 18-carbon aliphatic acids on the floatability of hematite, goethite, and of activated quartz. Electrophoretic mobility measurements, simplified flotation tests using a modified Hallimond tube, and contact angle and frothability measurements were employed. The results are of direct application in the interpretation of the batch flotation results on natural ores as

Jan 1, 1961

By H. A. Domian, H. Biloni, G. F. Bolling

Jan 1, 1965

By J. P. Carson

The works of the Tula Iron Company are in the Republic of Mexico, State of Jalisco, twenty-eight leagues southwest of Guadalajara, ten leagues northwest of the town of Sayula, through which passes the projected line of the Mexican National Railroad. Its geographical position would be about: latitude, 20' 10' N., longitude, 40 35' w. of Mexico City. The surrounding country is a rolling plateau, 6000 feet above the sea, enjoying the most magnificent climate in the world, the average temperature being about '70'. The works were commenced in 1850 by a company with very small capital, having not the least idea of the undertaking. They soon fell into the hands of the money-lenders, and after changing

Jan 1, 1879

By L. F. Pease, J. H. Brophy

A phase diagram for the Ta-Zr system is presented. The system is of the minimum-melting point type with the 0-zirconium phase decomposing monotectoidally at 785°C and 95.5 at. pct Zr. The minimum solidus temperature is 1875°C and the temperature of the top of the monotectoid loop is 1775°C. Techniques employed included X-ray diffraction and fluorescmce, electron microprobe analyses, diffusion couples, metallography, and melting-point measurements. Errors due to interstitial solutes and in melting observations are particularly common in this system, and their occurrence is discussed. SEVERAL versions of the Ta-Zr phase diagram have been reported in the literature.lm7 Williams et al.' report a diagram of the minimum-melting point type with a monotectoid reaction at 800°C and 12 wt pct Ta. The minimum in the solidus occurs at 1820°C and 25 wt pct Ta, while the top of the monotectoid gap is located at 1780°C and 80 wt pct Ta. The experimental techniques employed were metallography, electrical resistance variation with temperature, dilatometry, X-ray diffraction, and direct observation of melting by the technique of Pirani and ~lterthum.' Experience in the Nb-Zr phase diagram has shown that small changes in the interstitial content can cause large shifts in the positions of the phase boundaries. Small amounts of inter -stitials can cause similar changes in the Ta-Zr system and may be responsible for the wide variations in the phase diagrams previously reported.'-' The authors15 have found that alloys containing 0.75 wt pct 0 and 1.0 wt pct Ni with tantalum contents from 19 to 85 at. pct often contain a lamellar or cellular precipitate after annealing above 1400°C. This precipitate is very similar to that reported by pollack,12 Ralls," and Donlevy17 for the Ni-Zr system. In the present investigation, a wider variety of experimental techniques and improved alloy purity were applied to the resolution of errors in diagrams for this system as they now appear in the literature. New results have been obtained in the areas of alloy homogeneity and purity. Errors in the common techniques of direct melting observation in binary alloys have been demonstrated. In the commercially important high-tantalum region, the more accurate metallographic method of solidus determination has been employed. EXPERIMENTAL PROCEDURE Nominal analyses of the materials employed are given in Table I. The zirconium was low hafnium reactor grade iodide crystal bar, and the tantalum was in the form of high-purity sheet. Powders were impossible to use because of gas content. Alloys were melted on a water-cooled copper hearth, with a thoriated tungsten-tipped nonconsumable electrode in 2/3 atm of Zr "gettered" argon. Alloys were

Jan 1, 1963

By A. W. Ruff

The seismic wave produced by an explosive is very important in blasting. A true understanding of the wave is only important when considering possible structural damage to buildings located near the blast, but it is also of primary importance when considering rock fragmentation. Because of the importance of the seismic wave to fragmentation, a finda-mental study of the seismic wave and its effect was made at the Cananea Consolidated Copper Co.'s mine at Cananea, Sonora, Mexico. Strain gages were used to measure the strain in the rock when nearby explosive charges were detonated. The variance in the circuit voltage caused by changes in the resistance of the strain gage was measured with an oscilloscope and recorded with a Polaroid camera. CONFINED DRILLHOLE CHARGES To record the seismic wave from a confined drillhole explosive charge, a gage was cemented to the bottom of a 4-in diam drillhole. The horizontal gage hole and nine horizontal charge holes were drilled perpendicular to the center line of the 2-39 drift and 4 ft above the floor. The gage hole was 29 in. deep, and the charge holes were drilled so that each bottomed 1 in. deeper than the gage hole. The gage was installed with its axis horizontal and parallel to the 2-39 drift. The charge holes had the following horizontal distances from the gage: hole 3, 3 ft; hole 4, 4 ft; hole 5, 5 ft; hole 9, 9 ft; hole 10, 10 ft; hole 14, 14 ft; and hole 15, 15 ft. Figs. 1 and 2 show the relationships of the gage and the charge holes. Charges of 80 pct special gelatin, which weighed 50, 100, and 150 g were exploded in the charge holes and the seismic waves were recorded. All charges were stemmed with 200 g of mud and detonated with an electric blasting cap. Fig. 3 shows the recorded seismic waves. The reflected components of the waves were too weak to be objectional on these records. Table I shows the data taken from the wave forms. The approximate average velocity of a wave with a 9-ft travel distance was 12,900 fps. Since different charge weights were used, it was necessary to scale the data. The scaled maximum compress ion-strain data and the scaled maximum tens ion-strain data are shown in Figs. 4 and 5, respectively. Figs. 6 and 7 show the strain-distance data from Figs. 4 and 5 converted to a 1-lb charge. The results showed that the wave consisted of an initial compression pulse followed by a tension pulse. The amplitude of the tension pulse approached the amplitude of the compression pulse as the distance between the gage and the shot increased. The decay rate of the seismic wave is very important in the evaluation of the blasting characteristics of a rock. When high-amplitude stresses are produced, as in rock fragmentation, the rate of decay is much more rapid than when small amplitude stresses are produced. In underground mine blasting, the burdens are usually not greater than 6 ft and the decay is very rapid. The rapid seismic-wave decay rate in the fragmentation range means that in practice, if the minimum weight of powder is used to break a specified burden, only a very

Jan 1, 1961

By Arthur L. Hall

The purpose of this symposium is to challenge industry to better air classification, and as a first step toward fulfillment, to spotlight three fairly recent, aerodynamically designed, highly efficient air classifiers. This discussion does not intend to imply that present practice is bad. Actually, when present air separators are used for the jobs for which they were originally designed, they give high production and uniform operation, with a minimum of maintenance. The customers of nonmetallic mineral pigment producers, however, are making increasing demands for finer products, such as 99.9 pct -20µ, which appears to be beyond the economic capabilities of standard air separators industrially used over the last 10 to 30 years. Here is an actual example. A 10-ft diam pebble mill is in closed circuit with a conventional 12-ft diam centrifugal air separator, grinding 20-mesh wollastonite to a series of fine grades up to 99.97 pct - U.S. No. 325 mesh (44µ). Fig. 1 shows variations in fineness of mill discharge (or air separator feed), recycle for regrinding (or air separator tails), and product as measured by a wet rotating sieve test. For example, when the mill is turning out a product 80 pct -325, the mill discharge or separator feed is 30 pct -325, and the coarse tailings or recycle is only 5 pct -325. Hence 25 pct of the 30 pct available is accepted as product. This is efficient separation, and the ratio of recycle to feed is only about 2: 1. How- ever, to make a product testing 99.97 pct -325 the separator feed is 62 pct -325 and the coarse tailing 60 pct -325. In this case the recycle ratio is about 13: 1 and the economic limit to produce fineness has been substantially reached with this given equipment and process arrangement.

Jan 10, 1957

By S. G. Gibbs, A. B. Neely

Jan 1, 1967

By J. Douglas, W. T. Ford, G. E. Henderson, I. F. Roebuck

An implicit numerical method is presented for simulating the differential and algebraic relations governing one-dimensional three-phase flow in porous media. The method is based upon compositional representation of the hydrocarbon system. Variable physical properties and water-oil capillary forces are included in the formulation; however, the effect of gravity is ignored. The effects of changing composition and mass transfer are considered through the use of known phase behavior concepts and correlations. The differential and algebraic equations and the numeric approximations to these equations are presented, and the computing algorithm is discussed in detail. This method is used to simulate a single-phase, two hydrocarbon-component gas displacement laboratory experiment. The method is also applied in simulating a solution gas drive and a gas injection problem using a reservoir oil represented as a nine-component mixture. For comparative purposes, the latter two problems also are examined with a one-dimensional volumetric model. The results of the simulation of the laboratory experiment demonstrate the relevance of this method to problems wherein the effects of mass transfer significantly enter the displacement process. The comparison of the results of the volumetric and compositional methods applied to a reservoir system illustrates not only the advantages and benefits derived from examining the hydrocarbon system composition, but also the necessity of this approach in application to a general reservoir problem. INTRODUCTION The advent of large memory, high-speed digital computers enabled reservoir engineers and mathematicians to pool their knowledge in the development of sophisticated techniques for the prediction of oil and gas reservoir performance. These techniques have progressed from volumetric and trend analysis calculations to the development of "mathematical models", or reservoir simulators. These latter models, in general, utilize finite difference approximations to the rather complex partial differential equations that mathematically describe the physics and thermodynamics of fluid flow in porous media. Many forms of this "model" type are available today, and much of the recent literature has been devoted to either improving known methods or developing new techniques for use in these prediction methods. The more advanced techniques in use at this time consist of the solution of differential and algebraic relations that describe a reservoir geometry in one, two or three dimensions and that incorporate reservoir hydrocarbon and water fluid systems that obey specified physical laws, having properties that might easily be measured in the laboratory. This type of model utilizes formation volume factors and other volumetric data, expressed directly as functions of pressure, to represent the reservoir fluid system. Experience gained with this type of simulator, commonly known as a "beta-factor model", has indicated that the approach has many outstanding defects, most of them due to the assumption that those properties measured in the laboratory completely describe the behavior of the fluid system throughout the life of the project under study. The effects of varying hydrocarbon composition and mass transfer between the phases of this hydrocarbon system must, by definition, be ignored. These defects, which may often be ignored in

Jan 1, 1970

By L. F. Elkins

In 1946 Littlefield, Gray and Godbold published a thorough geologic description of the West Edmond Hunton Lime reservoir, located in Central Oklahoma, and discussion of its early performance.l They stressed implications of the nature of the reservoir rock on performance to be expected with production both by natural means and by pressure maintenance with gas injection. The essence of their analysis was that presence of an extensive interconnected system of fractures and solution channels constituting about 10 percent of the total reservoir void space in otherwise very low permeability reservoir rock would result in severe channeling of naturally encroaching water or injected gas through the fractures with little or no benefit to ultimate recovery of oil. Considerations were in progress to unitize the reservoir for pressure maintenance by gas injection. Some other engineers involved in the project had considerably different opinions at the time regarding continuity of fractures and regarding effects of such fractures on performance. The field was unitized Oct. 1, 1947. Subsequently, pilot gas injection tests were conducted in four areas of the field; large scale waterflooding was conducted in two parts of the field in the Bois d'Arc or Upper Hunton; and pilot waterflooding was conducted in the Chimney Hill or Lower Hunton. Full scale water-flooding is now in progress in the Chimney Hill section where it is oil productive. The primary purpose of this paper is to divine some of the secrets of this reservoir through analysis of all aspects of its performance, with the hope that it will be of help to engineers and operators in developing and operating other low permeability fractured reservoirs. In part, this considers variance of actual performance from some widely used conventional reservoir engineering assumptions. For brevity, other aspects of unit operation and its benefits to the owners are not discussed. Oil and Gas in Place For unit participation, net pay in the Bois #Arc defined by the gross SP anomaly on electric logs totalled 2,100,000 acre-ft in the 30,000-acre reservoir. Development of West Edmond predated modem logging techniques that now would be used to determine "net" or "productive" pay. In Fig. 1 this log pick is compared with detailed core analysis and core descripfion by Littlefield et a1.l for one well identified as A on the index map in Fig. 2. Not all of the gross Bois #Arc section had oil staining, but all cores with measureable intergranular permeability were stained. (The committee selection of the Bois d'Arc SP interval also includes the Frisco section as identified from cores by Littlefield.) The Bois d'Arc is oil productive throughout the entire 600-ft monoclinal accumulation from the eastern upstructure erosional pinchout to the west-side oil-water contact. (See Littlefield et al.l and McGee and Jenkins for geologic detail.)

Jan 1, 1970

By J. W. MacDonald

Review of equipment, performance, and shaft design plans including head-frames for sinking and forms for lining are given. Also included are comparative data for shafts of 11 x20 ft and 193/4 x 30 ft with 756 and 666-ft depths, respectively. Advantage from curtain wall installation with "slip forms" which warrant consideration for normal lining are outlined. A preliminary review is also made of "chemical grouting" to stop water inflow leakage with Halliburton equipment and service. Two coal mine shafts recently completed in Franklin County, Ill., afford experience and accomplishment to warrant review. These shafts provide access to a tract with a potential recovery of one hundred million tons. This work provided the initial sinking by other than a contractor's organization since 1918. Four additional shafts required during the interim were sunk under contract arrangement with others. Supervisory contact and inspection afforded experience to govern planning and conduct of the recent work. TEST DRILLING Core test holes were drilled within each of the two shaft locations. Four additional cable tool holes were drilled on a radius of 700 ft from the shaft bottom to determine the coal seam elevations and govern the underground layout plans for each shaft. A graphic log of the hoist shaft test drillhole is shown by Fig. 1. A very soft sandstone was found at a depth of 16 ft. Cored recovery of the sandstone proved much weaker than the overlying compacted sand secured with coring facilities for soil recovery. Gradation from the very soft sandstone to fairly hard material occurred with the next 14 ft, extending to a depth of 30 ft. The weak character within the upper portion entailed need of adequate protection when sinking as described later. A part of the sandstone provided sufficient porosity and permeability to warrant further investigation. Adjacent cable tool drilling for temporary water supply disclosed inflow was limited to 2 gpm with a depth of 40 ft. It was increased to 12 gpm during the next 50 ft of drilling and occasioned subsequent precaution when sinking. General arrangement of facilities for sinking the hoist shaft is shown by Fig. 2. An available steel frame building was used to accommodate the hoist, compressor, shop, and shower or change room. The stiff leg derrick with 90-ft length of boom permitted removal of forms from the wood shop to the storage area or delivery within the shaft. Reinforcing steel and other needs were similarly handled, in addition to headframe erection. The headframe was placed over one end of the shaft to enable passage of forms for lining through the adjacent half next to the stiff leg derrick. The steel headframe as shown by Fig. 3 was fabricated locally. Side anchored cables from the

Jan 1, 1961