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By M. Albertson

The purposes of this statement are to define a problem that exists in regard to the estimation of developed petroleum reserves, to analyze the problem in an abstract manner, and to discuss it as an introduction to certain papers that follow. The problem is to devise and develop a suitable procedure to permit more accurate estimates of developed petroleum reserves. Part of the problem is to make such estimates available earlier in the life of any particular pool. The problem of securing more accurate estimates of developed petroleum reserves existed prior to curtailment of production. With curtailment, however, the problem became acute, some of the more prominent reasons being as follows: 1. The standard method of estimation depended on production-decline curves. This method is inapplicable while curtailment prevails and available methods arc insufficient. 2. The retarding influence imposed by wide open flow conditions upon overdevelopment is absent. Great losses of capital by overdevelopment can now occur before realization of the true situation occurs, or before effective steps can be taken to stop such wastage. This increases the need for early, accurate estimates. 3. Above-ground stocks have diminished and the equivalent of such supply tends more and more to be retained underground in the natural reservoirs. It becomes increasingly necessary to know accurately the amount of these stocks. To direct attention to an additional point of different nature: improved accuracy of reserve estimation can be obtained only by closer scrutiny of the factors that determine recovery. This examination and the resulting efforts to measure the effect of these factors in terms of ultimate recovery must tend to improve ultimate recovery. The statements regarding the acute nature of the problem also indicate to some extent the importance of the problem and the necessity of finding a solution. Developed reserves of the industry, and of individual opera-tors, are one of the great factors that properly should determine explora-

Jan 1, 1936

By A. T. Peters

DURING an investigation concerning the use of large nozzles for pouring of low-carbon rimming steels, the practical details of which were reported elsewhere, it was noticed that increases of nozzle sizes were generally followed by improvements of slab surface quality. It is the purpose of this note to present the data available and some tentative opinions about this phenomenon. The problem of correlating pouring rates with the metallurgical performance of steels is complicated by several factors. The two most important are; 1) Variations of the steel pouring rate throughout a pour, these being due to the nozzle erosion, which enlarges the bore area, and the dropping ferrostatic pressure in the ladle. The pouring rate is lowest towards the end of a heat when the ladle is nearly empty. 2) Variations in the ingot pouring time due to the different ingot weights poured. Consequently, the use of the average pouring rate of a heat may not be exact, but in large-scale production it is not practical to treat ingots separately according to their pouring rates. The variations of ingot weights and sizes may be overcome by considering the rate of rise of steel in the mold. Hence, the average rate of filling of molds throughout the pour of a heat, inches per second, will be used in this note. Every ingot rolled at Inland is rated at the blooming mills prior to hot machine scarfing as being good, fair, or poor for cracks, scabs, and checking. The latter term refers to the network of breaks or lines associated with overheating or burning in the soaking pits. It is also known that thin-skinned steel is prone to show this defect if the subsurface blowholes are exposed during reheating or rolling. The experiences with the surface quality of slabs rolled from ingots poured at different rates are presented in Fig. 1. The data represent nearly 11,000 ingots. It appears that all three types of defects are in-

Jan 1, 1960

By A. B. Kinzel

SINCE the exposition of the behavior of certain iron alloys by Sykes1 involving the existence of an austenite loop and the discovery of such a loop in the chrome iron system by Bain,2 there has been much conjecture both on his part and on the part of metallurgists in general as to the exact shape and location of this loop. In previous work commercial steel has been used and the microscope has been employed to locate the critical points so that the position of the loop has been quite approximate. Numerous attempts have been made to measure the critical points of the carbon-free alloys but all of these for one reason or another have failed. A satisfactory method for studying these phenomena is presented here. MATERIAL The first difficulty in this study was to obtain pure chromium-iron alloys. This has been overcome by melting in a vacuum electrolytic Eton and chromium analyzing 99.84 Cr, Fe and C. A pure magnesite crucible was used and the melt allowed to solidify in the crucible. A typical analysis is chromium, 10.10 per cent., carbon, 0.006 per cent., silicon and manganese nil. To overcome the problem presented by segregation, ingots were made into wire 0.04 in, in diameter. This was accomplished by hand-forging the ingots into rods, using an oxyhydrogen flame for heating, in order to avoid carbon. Subsequently the rods were swaged to wire, the heating being carried out in hydrogen. Thus local segregation was eliminated although differences in analysis between portions 50 ft. apart were noted. An electric-resistance method was first tried to determine the critical points, but due to the physical defects in the wire this failed. Bain had previously shown the presence or absence of critical points in a qualitative way by observing the sag in a catenary. The author has devised the following method which gives an accurate determination of the critical points from expansion-temperature curves of suspended wire, a special telescopic dilatometer being used for the purpose.

Jan 1, 1928

INTRODUCTION The mineral sample is the fundamental unit upon which the mineral appraisal is built. Values derived from the sample are the most important quantitative measurements made in the total evaluation process. The sample value

Jan 1, 1980

By W. G. Pfann

WHEN a molten zone traverses a long, solid ingot in a level, open boat the ingot becomes tapered. While the taper may be slight after one zone pass, it can be appreciable after repeated zone passes, even to the extent of causing an overflow of matter at the end of the boat. It will be shown here that such matter transport by a molten zone arises from the volume change on melting and that it corresponds in magnitude and direction to the magnitude and sign of the volume change. It will be shown also that the mechanism of matter transport is basically similar to that of solute transport by a molten zone, which has been discussed elsewhere,1,2 and that the two processes can be described by similar equations. Finally it will be shown that matter transport can be prevented by inclining the ingot at a critical angle. Contraction on melting results in forward transport, that is, transport in the direction of zone travel; expansion on melting results in reverse transport. The way in which forward transport arises from the contraction on melting may be seen as follows: At the starting end of a solid horizontal ingot of uniform rectangular cross section of height h., form a molten zone of length 1, as shown in Fig. 1. The height h of the liquid will be h. where a is the ratio of density of solid to that of liquid. Advance the molten zone, at constant length, a distance dx. The volume of solid melted will be hdx; the volume of solid frozen will be Izdx, because the height of the freezing solid will tend to be equal to that of the liquid. The resultant change, dv, in liquid volume will be the difference in these quantities multiplied by the density ratio. Hence the following equation may be used for a zone at any distance x from the origin: which is valid in all but the last zone length in the

Jan 1, 1954

By Kenneth Cottingham

The area described herein is situated on what is known as the Arches Fork anticline in Roane and Calhoun Counties, W. Va. When the area was first mapped, it was felt that well records would give a more intimate idea of conditions, both structural and lithologic, than could be expected from surface horizons. The surface formations are chiefly members of the Monongahela series, though northward the hilltop outcrops are those of the Dunkard. Neither series offen a key horizon sufficiently definite to be satisfactory, although the structure as determined from well records was later checked against two of the more persistent surface sandstones. In mapping the subsurface structure, the top of the Big Lime (Green-brier) was employed as the key horizon. The average depth to the top of the Big Lime varies from 1600 to 2000 ft. (487 to 609 m.), while the average thickness as indicated by well records is 113 ft. (34 m.). The oil-producing formation is the Big Injun, lying just below the Big Lime. In two instances, well records show the Lime and Injun separated by a few feet of "break." The Big Injun is recorded as having a variable thickness. Seven logs report it as missing and one gives it the maximum thickness of 105 ft. (32 m.), while the entire average for thickness is 19 ft. (6 m.). Relation of Structure to Thickness of Big Injun The seven wells reporting the Big Injun as absent were located near the anticlinal axis. Moreover, the general tendency is a thickening of the Injun on the flanks of the anticline, increasing with distance from the axis. Manifestly, this apparent anomaly has no stratigraphic basis, as the actual thickness of the Big Injun and the overlying Big Lime was determined by original deposition which no amount of subsequent folding could appreciably alter. Fig. 1 is a skeleton map of structure based on the top of the Big Lime; Fig. 2 is a cross-section showing the thickness of the Big Injun.

Jan 1, 1923

By Kenneth Cottingham

The area described herein is situated on what is known as the Arches Fork anticline in Roane and Calhoun Counties, W. Va. When the area was first mapped, it was felt that well records would give a more intimate idea of conditions, both structural and lithologic, than could be expected from surface horizons. The surface formations are chiefly members of the Monongahela series, though northward the hilltop outcrops are those of the Dunkard. Neither series offen a key horizon sufficiently definite to be satisfactory, although the structure as determined from well records was later checked against two of the more persistent surface sandstones. In mapping the subsurface structure, the top of the Big Lime (Green-brier) was employed as the key horizon. The average depth to the top of the Big Lime varies from 1600 to 2000 ft. (487 to 609 m.), while the average thickness as indicated by well records is 113 ft. (34 m.). The oil-producing formation is the Big Injun, lying just below the Big Lime. In two instances, well records show the Lime and Injun separated by a few feet of "break." The Big Injun is recorded as having a variable thickness. Seven logs report it as missing and one gives it the maximum thickness of 105 ft. (32 m.), while the entire average for thickness is 19 ft. (6 m.). Relation of Structure to Thickness of Big Injun The seven wells reporting the Big Injun as absent were located near the anticlinal axis. Moreover, the general tendency is a thickening of the Injun on the flanks of the anticline, increasing with distance from the axis. Manifestly, this apparent anomaly has no stratigraphic basis, as the actual thickness of the Big Injun and the overlying Big Lime was determined by original deposition which no amount of subsequent folding could appreciably alter. Fig. 1 is a skeleton map of structure based on the top of the Big Lime; Fig. 2 is a cross-section showing the thickness of the Big Injun.

Jan 1, 1923

By Howard Willey

Mining operations of the Charcas unit at present are limited to the Tiro General mine at Charcas, in the State of san Luis Potosi, Mexico. The Tiro General mine was first operated during the Spanish occupation of Mexico for rich silver ores in an oxidized zone along the outcrop of the principal vein. In depth the ore is a complex sulfide with widely varying proportions of sphalerite, argentiferous galena and chalcopyrite ; and until recent years, it was economically impossible to carry on mining otherwise than irregularly in sections of the veins richest in galena and chalcopyrite. The advent of selective flotation brought the orebody as a whole within the field of economic interest, and a flotation mill of 650 metric tons daily capacity was constructed in the year 1925. The mining of partly worked-out sections of the mine, as well as the very heavy ground in most of the stopes opened in virgin ore, presented an unusually difficult problem. This paper describes the so-called "filled top slice" stoping method finally adopted as the standard system. Physical and Geological Conditions The principal orebody of the mine is an oreshoot about 1400 ft ng in a fissure vein striking about east and west and with an average lip of 70" to the north. The footwall of the orebody is fairly cleancu and regular both longitudinally and vertically and generally shows a dell-marked post-mineral slip. The footwall rock varies from a fairly strong limestone to a broken, metamorphosed limestone which requires heavy timbering. In some sections it is an altered, crumbly porphyry, which swells badly and necessitates frequent renewal of drift and raise tirnber. The hanging wall is irregular and frequently an assay wall. The stopina width ranges from a few feet up to more than 90 ft.; it averages about 30 ft. For the most part the hanging-wall rock is a shattered and altered monzonite porphyry, nearly aways requiring heavy timbering. The orebody is greatly fractured by post-mineral movement, which usually followed the footwall and left it as a smooth surface frequently covered with a slippery clay gouge. Two well marked transverse faults

Jan 1, 1931

By C. R. McEwen

This paper presents a technique for calculating the original amount of hydrocarbon in place in a petroleum reservoir, and for determining the constants characterizing the aquifer performance, based on pressure-production data. A method for doing this based on a least-squares line-fitting computationwus proposed by vanEverdingen,Timmerman andMcMahon in 1953. We found that their method would not work when there is error in the reservoir pressure data — even moderate error. The technique presented here appears to give reasonable answers when pressure data are uncertain to the degree expected in reservoir pressure determinations. The major change introduced in the present analysis is to limit the least-squares line-fitting to yield only one constant— the amount of hydrocarbon in place. The water-influx constant is then taken as proportional to the oil (or gas) in place. The constant of proportionality can be computed from estimates of effective compressibility and reservoir water saturation. It is also pointed out that the commonly used least-squares analysis assumes all of the uncertainty to be in the dependent variable. The material balance should be arranged so that this condition is fulfilled if correct inferences are to be made from statistical calculations. Examples are shown of the application of the new technique to gas reservoirs—both hypothetical and real—and to the oil-reservoir example of van Ever-dingen, Timmerman and McMahon. INTRODUCTION The amount of hydrocarbon originally in place in a petroleum resenroir can be estimated by means of the material-balance calculation. Simultaneous observations of pressure and amounts of produced fluids are required, together with the PVT data applicable to the reservoir fluids. If water encroachment is occurring, it is desirable to try to infer the behavior of the aquifer, as well as the original hydrocarbon in place, from the pressure-production data. This imposes additional demands on the method of calculation, and uncertainty in the data can result in large uncertainty in the answer. In addition, if the size of a gas cap is to be established, the whole problem becomes indeterminate, as pointed out by Woods and Muskat.1 Brownscombe and Collins2 simulated a gas reservoir and its aquifer on a reservoir analyzer and derived quantitative information on the effect of uncertainty in pressure and aquifer permeability on computed gas in place. Among the various techniques of estimating the performance of an aquifer, the method of van Ever-dingen and Hurst? based on compressible flow theory, seems to have been the most generally successful (see Ref. 4, for example). In this paper we shall confine ourselves to their representation of the aquifer. In 1953,' van Everdingen, Timmerman and McMahon5 introduced a statistical technique for deriving the amount of oil originally in place and the parameters which describe the aquifer. (We shall refer to this technique as the "VTM method'', as suggested by Mueller.6) Their example reservoir had no gas cap. It has been our experience that the VTM method gives a reasonable answer when the data are very accurate, but that inaccuracy (particularly in pressure) can cause the method to break down. The effect was first observed in gas reservoirs, but has since been seen in oil reservoirs also. In this paper we present another statistical method which has been successful in achieving a reasonable answer where the VTM method has failed. In the new method, one less parameter is derived from material-balance computations. It is assumed that values can be established for effective compressibility in the aquifer and reservoir water saturation independently of the material-balance calculation. The water-influx constant can then be obtained from these data and the quantity of hydrocarbon in place.

By William F. Blake

The discovery of tin-ore in the Black Hills of Dakota dates from June, 1883, and short preliminary notices were published in September of that year.* The discovery-point is known as the Etta Mine, in the central portion of the Hills, about six miles east of

Jan 1, 1885

By William H. Callahan

"For his exceptionally successful career as a geological mentor, for his ore-finding skills, and particularly the discovery of the Central Tennessee Zinc District and for his lecture, 'Prospecting in the USA-Promise and Problems.' " Recipient of the 1982 D.C. Jackling Award

Jan 1, 1983

By Clement K. Chase

The ammonia leach for copper recovery is discussed from historical and current practice standpoints. The chemistry of the system is presented together with Eh-pH and copper-ammonia solubility curves. A list of metals forming ammine complexes is presented. The advantages and disadvantages of the ammonia leach as applied to copper recovery in conjuction with modern liquid ion exchange techniques are compared.

Jan 1, 1980

By L. G. Truby

This paper presents the results obtained after calculating matches of the observed pressure performance of five fields completed in a common aquifer. A general dejcription of the Central Basin Platform area in West Texas in which the five fields, Andector, Embear, Martin, TXL, and wheeler, are lo. cated is contained in the paper. The method of utilizing the electric analyzer to calculate simultaneously matches of the observed pressure performance of the five fields is outlined. The determination of boundaries of pressure communication is discussed and the extent of pressure interference between fields consistent with the configuration of the area aquifer is shown graphically. INTRODUCTION Anomalies in the pressure performance of the Andeector Ellenhurger Field resulted in an investigation of pressure interference between several Ellenburger fields located on the Central Basin Platform in West Texas. Four fields, Embar, Martin, TXL, and Wheeler, were found to contribute significantly to the pressure drawdown at Andector. An initial investigation of the pressure performance of the Andector Ellenburger Field revealed that the reservoir pressure could not be matched by the calculation technique usually applied to water drive reservoirs. In the past, many have advanced the opinion that fracturing in the Ellenburger was a local condition restricted to the immediate uplifted area around the field, in which case each field would in all probability be surrounded by its own relatively small aquifer. Following several unsuccessful attempts to account for the pressure performance at Andector, a system assuming a common aquifer involving several fields in the immediate vicinity of Andector was placed on the electric analyzer and the effect of lo. interference was noted. A refinement of the initial analyzer calculations involved the determination of the approximate boundaries of the aquifer, the calculation of detailed matches of the pressure performance of the interfering fields, and an evaluation of the degree of communication between fields. DISCUSSION Geology and Development The fields considered in this work are located in the Central Basin Platform, a north-south trending structural feature in West Texas lying between the Delaware Basin to the west and the Midland Basin to the east. The locations of various individual Ellenburger fields in this general are shown in Fig. 1 with the five major fields which appear to be definitely intercommunicating colored in black and located in the circled portion of the figure. Production from the five fields was first obtained from the Embar Field discovered in July 1942, followed by the Wheeler Field in 1943, the TXL Field in 1945, and the Martin and Andector Fields in 1946. The total reservoir fluid production from these fields is indicated in Fig. 2. As illllstrated in this figure, very little production was taken from these fields prior

Jan 1, 1952

By J. J. Strohl, H. J. Schwellenbach

This paper relates the results obtained by decreasing the size of feed to a rod mill screen circuit. As would be expected, added production was gained and a finer grind with no tonnage loss was made possible. THE MacIntyre mill of the National Lead Co. has four parallel sections in which magnetic and gravity separations are made. Ore from a 5000-ton bin is fed by Hardinge constant weight feeders to a 6x12-ft open-end Marcy rod mill in each section. Rod mill discharge is bucket elevated to two 4x10-ft Tyrock screens whose oversize is returned to the rod mill. Screen undersize is pumped to Crockett separators for magnetite removal, and the Crockett tailings are sized and tabled to make an ilmenite concentrate. Minus 200-mesh material, overflowed by the Crockett and Dorrco sizers is thickened, de-slimed, and concentrated by flotation.' By 1948, the development of a successful slime flotation process and numerous improvements in the gravity circuit of the mill had increased the capacity of the available concentrating equipment. At the same time, improvements in the operating techniques of the dry crushing plant and improved feeding of the crusher by the mining department had appreciably increased the production rate per operating hour in that plant. Therefore, it was decided to take advantage of the added crusher capacity to produce a finer rod mill feed and in that way increase grinding capacity. At approximately the same time, finer screens were installed in the rod mill circuits to give finer grind and thus improve liberation of the magnetite and ilmenite from the gangue and from one another. Prior to the first part of 1948, a 3/4-in. limiting screen was used for sizing rod mill feed. The first reduction in crushing plant screen size was to 1/2 in. With this screen, the circulating load in the tertiary crushing circuit was too high and plant capacity dropped below a practical hourly rate. A 9/16-in.-sq opening screen, installed next, proved satisfac-

Jan 1, 1951

By H. J. Schwellenbach, J. J. Strohl

This paper relates the results obtained by decreasing the size of feed to a rod mill screen circuit. As would be expected, added production was gained and a finer grind with no tonnage loss was made possible. THE MacIntyre mill of the National Lead Co. has four parallel sections in which magnetic and gravity separations are made. Ore from a 5000-ton bin is fed by Hardinge constant weight feeders to a 6x12-ft open-end Marcy rod mill in each section. Rod mill discharge is bucket elevated to two 4x10-ft Tyrock screens whose oversize is returned to the rod mill. Screen undersize is pumped to Crockett separators for magnetite removal, and the Crockett tailings are sized and tabled to make an ilmenite concentrate. Minus 200-mesh material, overflowed by the Crockett and Dorrco sizers is thickened, de-slimed, and concentrated by flotation.' By 1948, the development of a successful slime flotation process and numerous improvements in the gravity circuit of the mill had increased the capacity of the available concentrating equipment. At the same time, improvements in the operating techniques of the dry crushing plant and improved feeding of the crusher by the mining department had appreciably increased the production rate per operating hour in that plant. Therefore, it was decided to take advantage of the added crusher capacity to produce a finer rod mill feed and in that way increase grinding capacity. At approximately the same time, finer screens were installed in the rod mill circuits to give finer grind and thus improve liberation of the magnetite and ilmenite from the gangue and from one another. Prior to the first part of 1948, a 3/4-in. limiting screen was used for sizing rod mill feed. The first reduction in crushing plant screen size was to 1/2 in. With this screen, the circulating load in the tertiary crushing circuit was too high and plant capacity dropped below a practical hourly rate. A 9/16-in.-sq opening screen, installed next, proved satisfac-

Jan 1, 1951

By Frank Ruskey, Richard G. Burdick

Many problems associated with potential hazards in a proposed mining area may not be adequately appraised because of the complexity of trying to evaluate them in terms of production requirements. Frequently, encountered hazards have to be dealt with by day-to-day operational schemes that require the miner to either force a way through them or avoid them by blocking off the area in which the hazards are known to exist. These considerations, after the mine is in full production, can be costly to the mine operator and can increase the danger to the miners. A hazard-evasion computer program for mine planning that will appraise proposed mine layouts on the basis of existing or potential hazards has therefore been developed. The program is intended for use during the planning phase of a mine, where it should assist the mining engineer in making decisions in mine design, With the output from this program, the mining engineer should be able either to corroborate his judgments made on the basis of mining experience or to find unanticipated solutions to design problems. The program tests hundreds of possible solutions to a mining layout, and by sorting out the 10 best (based on a criteria of minimum-abandoned-coal) limits the number of choices that need to be considered and appraised for applicability to a given mine scheme.

Jan 1, 1976

By Eizo Yabuuchi, Yukito Imanaga

INTRODUCTION In treatment of mine drainage, it is well known that the neutralization by calcium carbonate is far better than by slaked lime because of cheaper cost and better precipitability of its deposit. However, in case of iron ion contained in the mine drainage consists of ferrous salt, the ferrous salt must be oxidized to ferric salt before neutralization, because the ferrous salt does not react with calcium carbonate. At Yanahara Mine, during the period of 1950 through 1956, ferrous ion in the mine drainage had been oxidized to ferric ion, simultaneously neutralized by adding calcium carbonate into the aerated drainage. During the period of 1961 through 1974, after passing through many kinds of treating methods, the neutralization had been carried out by addition of calcium carbonate into the drainage after oxidized its ferrous ion to ferric by using nitrous gas as an oxidizing catalyzer. Meanwhile we had made an experiment on the oxidation of ferrous ion by utilizing iron bacteria, and developed a method which oxidized ferrous ion in large quantities briefly by cultivating and concentrating iron-oxidizing bacteria. Since 1974, we have been operating successfully for treating the drainage at Yanahara Mine by means of utilizing these bacteria.

Jan 1, 1976

By John V. Beall

Probably the only hotel in the world with near 100% occupancy at all times is the modern Hotel Indonesia in Djakarta. However, there is little gaiety among the scores of business men who occupy the hotel because they fret as days grow into weeks in seemingly interminable negotiations with government officials. As one man put it, "time is a dimension that they [Indonesians] have never grasped." But the visitors still come in droves to investigate business opportunities in this sprawling nation of tropical islands which is, with 120 million people, the fifth largest nation in the world. Approximately $300 million is being spent each year in Indonesia by foreign companies and governments. Aid grants from the latter are derived about one third from the United States, one third from Japan and the remainder from many nations. Djakarta is a far cry from pleasant colonial Batavia, as the city was called under the Dutch. The orderly colony of prewar days was managed as an agricultural and trading dependency of Holland. The Dutch did little about educating the people. The islands were controlled by the Japanese during World War II and afterwards there was a war for independence--which was attained in 1949. In the ensuing years, President Sukarno devastated the economy by squandering money on a gigantic scale. A Communist-attempted coup in 1965 resulted in a military takeover and the extermination of Communists. The new government under Suharto, which took over all power in March 1966, found itself in a situation where bare survival was the first consideration.

Jan 1, 1970

By Richard D. Call

Predicting the performance of a waste embankment requires an estimate of the properties of the waste and foundation materials. The performance characteristics and the associated material properties are: [ ] Methods for obtaining estimates of the material properties include: laboratory and field testing, back analysis, and indirect estimates from other material properties. The most critical technical need is the development of indirect methods for use in feasibility studies and mine planning, prior to production when actual samples of waste material are not available. Material properties are not unique, single values; they are probability distributions. There are two components to these distributions: the uncertainties in testing and the inherent variability of the material. The objectives of a material properties evaluation should be to obtain an unbiased estimate of the property distributions rather than the average or the minimum value. Whether the current trend to a probability design analysis is followed or a simple sensitivity analysis is conducted, the unbiased probability distribution is preferable to single values.

Jan 1, 1985

Jan 1, 1907