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OFFICERS AND DIRECTORS President and Director HOWARD C PYLE, '61, Los Angeles, Calif Past President and Director AUGUSTUS B KINZEL, '60, New York City President-elect and Director JOSEPH L GILLSON, '62, Wilmington, Del

Jan 1, 1959

By Sir Robert A. Hadfield

The Hadfield method of producing sound steel ingots has been the subject of a paper read before the Iron and Steel Institute, so that it will be unnecessary to describe it fully here. The object of the method is to make poseible the production of steel ingots in which the pipe is almost entirely obliterated, and it is applicable particularly to the production of sound ingots, from which good rails could be rolled. The method in brief consists in casting the ingots on cars, usually with the small end down. Each ingot mold is provided with a sand top, rammed up in the well-known manner. After the mold is filled to the desired height with steel, a layer of slag about 0.5 in. thick is placed upon it and the balance of the mold is then filled with charcoal. Then, through suitable piping, an air blast is directed in numerous jets upon the charcoal, which is burned thereby, the combustion supplying addi, tional heat to the top of the ingot, which helps to keep the top fluid and to retard its solidification, while the lower parts are rapidly losing heat by its transfer to the mold. By one plan the ingots are cast in molds placed upon bogies, movable trucks, or platforms, in the position A, Fig. 1. The ingots are cast vertically on bogies, say four or more ingots on each bogie, as may be found most convenient, and the upper surface of the steel is maintained in a molten condition during the transit of the ingots from the casting pit to the stripper. This is effected by a series of air pipes, for example, three in number, with jets or holes on the under surface of such pipes, placed preferably horizontally, arranged in such a manner as to cause jets of air to impinge on the upper surface of the ingots. These holes or jets are made throughout the entire length of the overhead pipes, the length of such pipes being

Jan 1, 1914

The program for the Annual Meeting to be held in New York, Feb.17 to 20, 1919, inclusive, has been nearly completed. Besides the usual meetings for the reading of technical papers, the social features are being given special attention. They will probably embody the following: On Monday evening, a lecture by one of our fighting airplane pilots, to be illustrated by lantern slides and moving pictures. To this entertainment the ladies will be invited, as well as members. On Tuesday evening, a Smoker at which some very startling features are promised for the entertainment of members. On Wednesday evening, the usual banquet, at which it is hoped to have some very interesting speakers. The usual luncheons will be served in the Engineering Building and other features will be offered for the entertainment of the ladies during the day. No definite plans for the Thursday excursion have yet been formulated, though it is probable that something interesting will be offered the members.

Jan 12, 1918

By Robert Bedford

THE objects of this paper are: 1. To give cost data on drilling obtained at the North Star mine in California; 2. To describe the methods of testing drills employed there, giving the results of shop tests, and the correlated data obtained underground; 3. To demonstrate that these tests form a satisfactory basis upon which to judge rock drills. The rock at this mine is either close-grained diabase in the upper levels, or tough grano-diorite in the lower levels. The vein, which has an average dip of 23°, consists of about 5 ft. of "formation" lying between walls of unaltered country rock. Of the 5 ft., solid quartz and stringers make up 18 in. of pay ore. The formation, which consists of but slightly altered country rock, is almost as hard as the unaltered walls. From 4 to 42 ft. is sent to the mill. One-third of the holes are drilled in the quartz; two-thirds in the formation. The holes are about 42 ft. deep and break on the average of 1.15 tons per hole. The average number of holes per stope drill shift is at present 5.65. The number of drill shifts throughout the mine in 1913, was 18,679, and the cost of labor for drilling, power, supplies, upkeep of machines and air lines, tool sharpening, and distribution, amounted to 33 per cent. of total mining expense. To do this work there are in commission 43 No. 12A, three No. 17V and eight No. 16V Waugh stopers, three Jackhamers, two Waugh pluggers, and 17 No. 8 ,and two No. 7 water Leyners. For the past three years a record has been kept of the repairs on each machine and number of shifts operated by it. The figures given in Tables I and II for the cost of repairs per drill shift have been taken from this record. The underground air pressure is about 90 lb. The air consumption per drill shift was measured by, aerometers. The leakage of pipe lines between compressor and meter is not included. From the number of pieces of steel sharpened for each type of drill during the year the average used per drill shift has been ascertained. For other supplies the figures given represent a year's average.

Jan 8, 1914

By John Mills

THREE case histories from professions other than engineering will serve to introduce ideas basic to this discussion. Case (1) Date, about 1900. A young man, B. D. from a three-year graduate course in a prominent divinity school, was for two years a successful pastor of a small church; then friction over theological matters arose with a fundamentalist deacon of the "Bible belt" type. He resigned, returned to his university for a four-year medical course, interned for two more, and thereafter was successful and apparently happy as a physician. We may assume that he was converted in early adolescence and vowed himself to service as a minister. Acceptance of his announced career by parents, friends, and teachers plus his own determination confirmed his choice until illiberal criticism threatened his usefulness and he awoke to the fact that his inner needs were not along the direction in which inertia and his conscience had constrained him.

Jan 1, 1947

By G. G. Sarkar

SHARPNESS of separation and washing efficiency are often confused. The sharpness index indicates the precision of a washing unit; the washing efficiency takes into account not only the unit's precision but also the washability characteristics of raw coal and the near-gravity materials at the effective density of cut. A sound basis for evaluating the sharpness index is essential in comparing the sharpness of separation attained by different cleaning units. Sharpness Index and Related Criteria: Washing plant sharpness of separation has been well represented by Hancock's efficiency chart and Tromp's curve of migrated materials. Some useful coefficients derived from Tromp's curve-the error area, probable error, and imperfection value-have been widely used to measure sharpness of separation. These are defined in a number of papers presented at the first and second International Coal Preparation Conference held in Paris (1950) and Essen (1954).1-5 In evaluating the performance of a dense medium separator at four different gravities, Yancey and Greer6 observed that even with similar operating conditions and the same size grading of feed, the imperfection value of a washing plant varies through a greater range than the probable error. The present writer also observed the same type of anomalies in comparing the imperfection values of one plant which were determined under almost identical conditions. He therefore devised a graphical method to evaluate a washing unit's sharpness of separation and define it by a single coefficient, which is substantially independent of coal characteristics under normal operation of the unit. Unlike imperfection value, the new coefficient can be expressed on a percentage basis.

Jan 12, 1957

By Abdul-Fattah K. Kaddou, P. C. Rosenthal

THE authors have studied the problem of temper brittleness in steel1 employing an internal friction method. The specimens are in the form of wire, 14 B and S gage (0.064 in diam), 6 in. long. It was very desirable to have an apparatus to test the wire specimens for temper brittleness without making standard notched-bar Izod or Charpy impact tests. Such a device would have two favorable advantages: the first is that no standard impact specimen would be necessary, thus saving time and labor; the second is that there would be a direct correlation with internal friction results, thus eliminating the size factor if standard impact specimens had to be used. A search was made without result to find a commercial apparatus for this purpose. The literature was surveyed to investigate if such a test had been utilized. No such work has been reported in the literature; however, zapffe2 has described an apparatus used for testing for hydrogen embrittlement in

Jan 1, 1961

By C. T. Butler, W. W. Kay, R. D. Boddorff, R. L Ash

DRILLING and blasting in anthracite open-pit mines is a continuous problem to contractors and explosive engineers because of the diverse conditions caused by the nature of the geological formations, the extensive mining of the portions of coal beds near the surface, and the proximity of many strip pits to populated areas. Pennsylvania anthracite occurs in four separate long and narrow fields totaling only 480 sq miles. The coal measures are rock strata and coal beds that are considerably folded and faulted. The crests of the anticlines are eroded extensively. The beds outcrop on the mountain sides and dip under the valleys. At first only the upper portions of the syn-clines could be stripped. Now stripping to increasingly greater depths is economically possible, as is indicated by the fact that the proportion of freshly mined anthracite produced by strip mining has increased from 3.7 pct of the total tonnage in 1930 to 29.6 pct in 1950. Much of the rock overlying the deeper beds now being stripped is so extensively broken that considerable difficulty is experienced in drilling satisfactory blast holes and in using explosives in such manner as to insure a uniformly broken material easily removed by the excavating machinery. Such breaking of rock strata has occurred because the bed now being stripped has been mined extensively in former years by underground methods, and tops of gangways and chambers have subsequently failed. Draglines are used to uncover coal where the overburden can be moved with little or no re-handling. These machines range in size from those having a 2 cu yd capacity bucket on a 60-ft boom to those handling a 25 cu yd bucket on a 200-ft boom. Draglines are also used to strip to the bottom of the coal basins if the depth and the distance between the crops are not too great. For this type of operation blast holes are drilled full depth to the bed. These holes are commonly 30 to 90 ft deep; however, in exceptional cases, holes may be as shallow as 12 ft or as deep as 130 ft. Drilling is normally done for blasts of 12,000 to 60,000 cu yd of overburden, 30,000 cu yd being considered an average blast if vibration is not the controlling factor. Where the stripping of wide basins or the exposure of a moderately pitching vein makes the use of draglines impractical, dipper front shovels equipped with 4 to 6 cu yd buckets load into trucks. Overburden is removed in benches of 25 to 30 ft with blast holes drilled 4 or 5 ft deeper than the planned floor of the bench. For shovels under 5 cu yd bucket capacity the volume blasted varies from 8000 to 12,000 cu yd, whereas a volume of 30,000 to 50,000 cu yd of overburden is frequently blasted at one time for the larger shovels where vibration is not an important factor. During the past decade the churn drill, generally the Model 42-T Bucyrus-Erie blast hole drill equipped for drilling 9-in. diam holes, has become the most common blast hole drilling machine. Electricity powers half the churn drills in use and is preferred on the large strippings where electric shovels are operated and the working area is concentrated. On these operations the cost of additional electricity for the drills is less than the cost of fuel to operate diesel units because of the existing large demand load of the excavating equipment. Moreover, electric motors start more easily in cold weather and generally are less expensive to maintain. Diesel driven units are employed where a higher degree of mobility is required. The average drilling speed is 8 ft per hr, although in softer rocks a rate of 15 ft per hr is attained. Where rock is hard and strata is badly broken, drill speeds may be less than 2 ft per hr. Low drilling production results under these circumstances when loose material falling from the upper portion of the drill holes causes drill stems to be jammed. Rock formations vary so greatly in the region that a 9-in. diam churn drill bit may become dull after drilling only 2 ft or may drill satisfactorily for 56 ft; however, an average of 35 ft is usual in sandstone of medium hardness. Dull bits are hoisted to flat bed trucks by the sand line of the drill and are usually sharpened in the contractor's bit shop adjacent to the job. Care is generally taken to cover the thread end of the bit with a cap. To facilitate handling of bits around the drill, a heavy thread protector having an eye top is becoming more popular than the flat-top rubber or metal cap furnished with new bits. The 9-in. diam blast holes for a 25 to 30 ft bench are normally on 18x18 ft to 20x20 ft spacings, depending on the character of the overburden, although in broken ground 15x18 ft centers may be used to obtain better breakage and a more even bottom for the bench. The patterns of holes for shots

Jan 1, 1953

By R. D. Boddorff, R. L. Ash, C. T. Butler, W. W. Kay

DRILLING and blasting in anthracite open-pit mines is a continuous problem to contractors and explosive engineers because of the diverse conditions caused by the nature of the geological formations, the extensive mining of the portions of coal beds near the surface, and the proximity of many strip pits to populated areas. Pennsylvania anthracite occurs in four separate long and narrow fields totaling only 480 sq miles. The coal measures are rock strata and coal beds that are considerably folded and faulted. The crests of the anticlines are eroded extensively. The beds outcrop on the mountain sides and dip under the valleys. At first only the upper portions of the syn-clines could be stripped. Now stripping to increasingly greater depths is economically possible, as is indicated by the fact that the proportion of freshly mined anthracite produced by strip mining has increased from 3.7 pct of the total tonnage in 1930 to 29.6 pct in 1950. Much of the rock overlying the deeper beds now being stripped is so extensively broken that considerable difficulty is experienced in drilling satisfactory blast holes and in using explosives in such manner as to insure a uniformly broken material easily removed by the excavating machinery. Such breaking of rock strata has occurred because the bed now being stripped has been mined extensively in former years by underground methods, and tops of gangways and chambers have subsequently failed. Draglines are used to uncover coal where the overburden can be moved with little or no re-handling. These machines range in size from those having a 2 cu yd capacity bucket on a 60-ft boom to those handling a 25 cu yd bucket on a 200-ft boom. Draglines are also used to strip to the bottom of the coal basins if the depth and the distance between the crops are not too great. For this type of operation blast holes are drilled full depth to the bed. These holes are commonly 30 to 90 ft deep; however, in exceptional cases, holes may be as shallow as 12 ft or as deep as 130 ft. Drilling is normally done for blasts of 12,000 to 60,000 cu yd of overburden, 30,000 cu yd being considered an average blast if vibration is not the controlling factor. Where the stripping of wide basins or the exposure of a moderately pitching vein makes the use of draglines impractical, dipper front shovels equipped with 4 to 6 cu yd buckets load into trucks. Overburden is removed in benches of 25 to 30 ft with blast holes drilled 4 or 5 ft deeper than the planned floor of the bench. For shovels under 5 cu yd bucket capacity the volume blasted varies from 8000 to 12,000 cu yd, whereas a volume of 30,000 to 50,000 cu yd of overburden is frequently blasted at one time for the larger shovels where vibration is not an important factor. During the past decade the churn drill, generally the Model 42-T Bucyrus-Erie blast hole drill equipped for drilling 9-in. diam holes, has become the most common blast hole drilling machine. Electricity powers half the churn drills in use and is preferred on the large strippings where electric shovels are operated and the working area is concentrated. On these operations the cost of additional electricity for the drills is less than the cost of fuel to operate diesel units because of the existing large demand load of the excavating equipment. Moreover, electric motors start more easily in cold weather and generally are less expensive to maintain. Diesel driven units are employed where a higher degree of mobility is required. The average drilling speed is 8 ft per hr, although in softer rocks a rate of 15 ft per hr is attained. Where rock is hard and strata is badly broken, drill speeds may be less than 2 ft per hr. Low drilling production results under these circumstances when loose material falling from the upper portion of the drill holes causes drill stems to be jammed. Rock formations vary so greatly in the region that a 9-in. diam churn drill bit may become dull after drilling only 2 ft or may drill satisfactorily for 56 ft; however, an average of 35 ft is usual in sandstone of medium hardness. Dull bits are hoisted to flat bed trucks by the sand line of the drill and are usually sharpened in the contractor's bit shop adjacent to the job. Care is generally taken to cover the thread end of the bit with a cap. To facilitate handling of bits around the drill, a heavy thread protector having an eye top is becoming more popular than the flat-top rubber or metal cap furnished with new bits. The 9-in. diam blast holes for a 25 to 30 ft bench are normally on 18x18 ft to 20x20 ft spacings, depending on the character of the overburden, although in broken ground 15x18 ft centers may be used to obtain better breakage and a more even bottom for the bench. The patterns of holes for shots

Jan 1, 1953

By R. D. Boddorff, R. L. Ash, C. T. Butler, W. W. Kay

DRILLING and blasting in anthracite open-pit mines is a continuous problem to contractors and explosive engineers because of the diverse conditions caused by the nature of the geological formations, the extensive mining of the portions of coal beds near the surface, and the proximity of many strip pits to populated areas. Pennsylvania anthracite occurs in four separate long and narrow fields totaling only 480 sq miles. The coal measures are rock strata and coal beds that are considerably folded and faulted. The crests of the anticlines are eroded extensively. The beds outcrop on the mountain sides and dip under the valleys. At first only the upper portions of the synclines could be stripped. Now stripping to increasingly greater depths is economically possible, as is indicated by the fact that the proportion of freshly mined anthracite produced by strip mining has increased from 3.7 pct of the total tonnage in 1930 to 29.6 pct in 1950. Much of the rock overlying the deeper beds now being stripped is so extensively broken that considerable difficulty is experienced in drilling satisfactory blast holes and in using explosives in such manner as to insure a uniformly broken material easily removed by the excavating machinery. Such breaking of rock strata has occurred because the bed now being stripped has been mined extensively in former years by underground methods, and tops of gangways and chambers have subsequently failed. Draglines are used to uncover coal where the overburden can be moved with little or no rehandling. These machines range in size from those having a 2 cu yd capacity bucket on a 60-ft boom to those handling a 25 cu yd bucket on a 200-ft boom. Draglines are also used to strip to the bottom of the coal basins if the depth and the distance between the crops are not too great. For this type of operation blast holes are drilled full depth to the bed. These holes are commonly 30 to 90 ft deep; however, in exceptional cases, holes may be as shallow as 12 ft or as deep as 130 ft. Drilling is normally done for blasts of 12,000 to 60,000 cu yd of overburden, 30,000 cu yd being considered an average blast if vibration is not the controlling factor. Where the stripping of wide basins or the exposure of a moderately pitching vein makes the use of draglines impractical, dipper front shovels equipped with 4 to 6 1/2 cu yd buckets load into trucks. Overburden is removed in benches of 25 to 30 ft with blast holes drilled 4 or 5 ft deeper than the planned floor of the bench. For shovels under 5 cu yd bucket capacity the volume blasted varies from 8000 to 12,000 cu yd, whereas a volume of 30,000 to 50,000 cu yd of overburden is frequently blasted at one time for the larger shovels where vibration is not an important factor. During the past decade the churn drill, generally the Model 42-T Bucyrus-Erie blast hole drill equipped for drilling 9-in. diam holes, has become the most common blast hole drilling machine. Electricity powers half the churn drills in use and is preferred on the large strippings where electric shovels are operated and the working area is concentrated. On these operations the cost of additional electricity for the drills is less than the cost of fuel to operate diesel units because of the existing large demand load of the excavating equipment. Moreover, electric motors start more easily in cold weather and generally are less expensive to maintain. Diesel driven units are employed where a higher degree of mobility. is required. The average drilling speed is 8 ft per hr, although in softer rocks a rate of 15 ft per hr is attained. Where rock is hard and strata is badly broken, drill speeds may ' be less than 2 ft per hr. Low drilling production results under these circumstances when loose material falling from the upper portion of the drill holes causes drill stems to be jammed. Rock formations vary so greatly in the region that a 9-in. diam churn drill bit may become dull after drilling only 2 ft or may drill satisfactorily for 56 ft; however, an average of 35 ft is usual in sandstone of medium hardness. Dull bits are hoisted to flat bed trucks by the sand line of the drill and are usually sharpened in the contractor's bit shop adjacent to the job. Care is generally taken to cover the thread end of the bit with a cap. To facilitate handling of bits around the drill, a heavy thread protector having an eye top is becoming more popular than the flat-top rubber or metal cap furnished with new bits. The 9-in. diam blast holes for a 25 to 30 ft bench are normally on 18x18 ft to 20x20 ft spacings, depending on the character of the overburden, although in broken ground 15x18 ft centers may be used to obtain better breakage and a more even bottom for the bench. The patterns of holes for shots

Jan 1, 1952

By A. Filyk

The current plant practice at Molycorp, Inc., Questa Division, a molybdenum producer in northern New Mexico, is described. An outline of crushing, grinding, flotation, drying, and tailings disposal is included, with emphasis on metallurgy.

Jan 1, 1976

By Ernest Kirkendall

The amount of research done in the last fen years on the subject of diffusion in solid metals is significant of the importance of this problem. To review the literature dealing with diffusion is unnecessary here. Deschl in 1912 reviewed the work that had been done on diffusion in solids and gave a bibliography of the literature published before that time. Mehl's paper in 19362 is an excellent review of diffusion in metals and contains a large bibliography. Of importance also for it? review and bibliography is an article by 14. I. Krynitsky.~ While considerable material lias appeared in the literature on diffusion in brass, the amount of information is not great compared to that pertaining to some of the other alloys. Most of the information published is of a qualitative nature; that which is quantitative varies considerably. Thus the quantitative determination of the rates of diffusion of copper and zinc in alpha brass by a new method appeared worth while. In the present investigation the rate of diffusion of zinc in approximately the same composition alpha brass has been determined :it three temperatures. Procedure The method used consisted of plating copper on a beta brass, heating it, to allow diffusion to take place, and measuring the concentration of copper by determining the lattice constant of the alloy. This method is unique in that it uses a beta brass of such a composition that at the tem perature at which diffusion takes place the beta brass is saturated wit11 copper. Thus, while the beta brass supplies thc zinc, no diffusion takes

Jan 1, 1939

By Ernest Kirkendall

The amount of research done in the last fen years on the subject of diffusion in solid metals is significant of the importance of this problem. To review the literature dealing with diffusion is unnecessary here. Deschl in 1912 reviewed the work that had been done on diffusion in solids and gave a bibliography of the literature published before that time. Mehl's paper in 19362 is an excellent review of diffusion in metals and contains a large bibliography. Of importance also for it? review and bibliography is an article by 14. I. Krynitsky.~ While considerable material lias appeared in the literature on diffusion in brass, the amount of information is not great compared to that pertaining to some of the other alloys. Most of the information published is of a qualitative nature; that which is quantitative varies considerably. Thus the quantitative determination of the rates of diffusion of copper and zinc in alpha brass by a new method appeared worth while. In the present investigation the rate of diffusion of zinc in approximately the same composition alpha brass has been determined :it three temperatures. Procedure The method used consisted of plating copper on a beta brass, heating it, to allow diffusion to take place, and measuring the concentration of copper by determining the lattice constant of the alloy. This method is unique in that it uses a beta brass of such a composition that at the tem perature at which diffusion takes place the beta brass is saturated wit11 copper. Thus, while the beta brass supplies thc zinc, no diffusion takes

Jan 1, 1939

By F. H. Labee

Questionnaires and special letters soliciting information were sent to a great many geologists, petroleum engineers, independent operators, and representatives of large companies in Arkansas, Louisiana, Texas and New Mexico. Replies to more than 50 per cent. of these letters were received. A survey of the facts and opinions presented in this correspondence will be made a little later, but first some of the general aspects of the problem will be reviewed. By "unit operation,"1 we mean the development and operation of a single entire pool, or reservoir, of petroleum by one management, on a systematic and scientific plan which is best calculated ultimately to extract from the pay sand and bring to the surface of the ground the largest possible amount of oil with the greatest possible total profit. No one, as far as the author has been able to ascertain, denies the advantages, both scientific and practical, of unit operation. All are agreed that it is sound in theory. It permits the proper distribution of wells, which should always be spaced on the ground according to a regular geometric plan with the intention of maintaining the same spacing all the way down and into the pay sand, or pay sands. This spacing will be different for different fields, depending on such variable factors as the lithologic character of the pay formation, the nature of the controlling structure, the depth of drilling, the viscosity of the oil, and so on. From the viewpoint of engineering practice, as applied to economical extraction of oil, there is no more important consideration than well spacing—well spacing in the pay formation, and not merely on the surface of the ground —for the method of well spacing has a direct bearing upon all such engineering practices as pressure control, repressuring by gas, the water drive, and SO on. H. C. Miller2 summarized the advantages of an ideal cooperative development plan as follows: "Among the more important advantages of

Jan 1, 1930

By Henry M. Howe

COLUMBIA UNIVERSITY The Equilibrium Position of A 1.—Some of the most important data on this subject are collected in Table I. Definition of Ae 1.—-Just as we call A 1 of rising temperature Ac 1, and that of falling temperature Ar 1, so we may give the symbol Ae 1 to the equilibrium position of A 1. Preliminary Assumptions.—In deciding what weight to assign the various data, and indeed what data are to count, three assumptions should be made. These are : 1. The corollary from the phase rule, that the temperature of Ae 1 is constant and independent of the carbon content, though, of course, it may be changed by the presence of other elements. 2. That because neither in heating nor in cooling can the transformatioil begin before the temperature reaches Ae 1, though it may be delayed by lag till after the temperature has passed beyond Ae 1, therefore Ae 1 should be at or below the observed beginning of Ac 1 and at or above the observed beginning of Ar 1, and, a fortiori, it should be at or below the observed maximum of Ac 1, and at or above the observed maximum of Ar 1. It may be necessary to take these maxima into consideration in view of the experimental difficulty of detecting these beginnings. 3. Because of the extraordinary degree of lag in the immediate neighborhood of Ae 1, as proved by Brayshaw, little weight should be attached even to the most conclusive evidence that the transformation has not occurred on reaching, or even staying at, a given temperature, because that failure to occur may always be due to lag. Evidence that the transformation has occurred, or has even begun, at a given temperature deserves incomparably greater weight. In other words, the temperature may pass beyond Ae 1 without inducing the ' transformation, but the transformation cannot be induced except by

Jan 1, 1914

By Douglas R. Cook

The discovery of the gold deposits of the Jerritt Canyon district, with proven recoverable reserves in excess of 67.6 Mg (2.4 million oz) of gold, has contributed greatly to the tremendous resurgence in exploration for disseminated gold deposits in Nevada and throughout the west. The Jerritt Canyon project, including the Bell mine, is a 70130 joint venture between Freeport Gold Co. and FMC Gold Corp., both of which are wholly owned affiliates of the respective parent companies, Freeport-McMoRan Inc. and FMC Corp. The Jerritt Canyon district is located in the center of the Independence Mountains in north-central Elko County, Nevada, approximately 48 km (30 miles) northeast of the Carlin gold mine and 68 km (42 miles) north of Elko. Although the discovery of a commercial gold mine at Jerritt resulted from Freeport's exploration activities, the initial identification of a gold prospect in this area was a consequence of a program by FMC in the search for antimony. Their program was, in part, based on data in the Nevada Bureau of Mines Bulletin 61, "Antimony Deposits in Nevada." Work on the property began in 197 1 with geologic mapping followed by sampling and geochemical analyses. FMC geologists, including Robert Hawkins, Russell Hayden, and Hal Hurst recognized the striking similarity of the area to the geologic environment of Newmont's nearby Carlin gold deposit and soon focused on the gold potential. Increases in the price of gold served to stimulate interest and support for the exploration program. Geochemical samples were collected and analyzed for gold and trace elements usually associated with gold. A strong gold anomaly was found on the North Fork of Jerritt Canyon, which was called the Alchem anomaly. Drill-testing of this anomaly in 1973 revealed significant grades and thicknesses of gold mineralization in the lower portion of the Roberts Mountain formation. This initial discovery and subsequent close-spaced drilling proved the existence of several small pods of low-grade, gold-bearing material which cropped out at their up-dip edges, causing the surface anomaly. The mineralization was very encouraging, but was not, by itself, of economic importance from the standpoint of either grade or tonnage. Freeport Exploration Co. established a district office in Reno in late 1974 with responsibilities for exploration of hard minerals in the Basin and Range Province. Emphasis of the program was directed to the discovery of precious metal deposits. Enfield Bell, as District Manager for this new office, was particularly interested in the Independence Mountains and when the opportunity for a joint venture with FMC became available in 1976, it was aggressively pursued. Enfield Bell and David Stevens evaluated the FMC data and with the support of management formulated a bid in competition with 23 other companies. Freeport was selected by FMC as partner and a joint venture agreement was successfully negotiated. In mid-summer of 1976, Freeport Exploration Co. began an expanded program of detailed mapping and geochemical sampling under the direction of Bell and Stevens that led to new interpretations of the structural and alteration patterns. Hawkins' contribution during this period must be recognized since he not only was one of the original geologists assigned to the antimony and subsequently the gold investigations, but shortly after the start of the joint venture exploration, he was hired by Freeport Exploration Co. as the project geologist for the venture. Hawkins had studied the central Independence Mountains for his thesis and his understanding of the regional geology provided an excellent background for the continued exploration of the gold de- posits. His interpretation of the structure in the vicinity of the Alchem anomaly was of significant help in the Freeport exploration program. Drilling, based on this new under- standing of the geology, revealed the edge of the Marlboro Canyon ore body, which is now being exploited in the Bell mine. Although a classic bullseye geochemical target led to the discovery of the Alchem mineralization, the bulk of the reserves known today in the Marlboro Canyon area lay

Jan 1, 1985

THE 145th* meeting of the American Institute of Mining and Metallurgical Engineers was held in New York, Feb 17-21, 1936 On the fifth day, Feb 21, an Institute party journeyed to Schenectady, N Y, and made there a tour of inspection of the "House of Magic" of the General Electric Co The meeting proper (Feb 17-20) consisted of the annual business meeting, 43 technical sessions at which 214 papers were presented, two regular sessions of the Board of Directors, two conferences of Section and Division Chairmen, the second of which was attended by the Board of Directors, a late afternoon session for Student Associates and members of Affiliated Student Societies and Student Chapters, which was addressed by President Buehler, Dr George Otis Smith and others, two formal lectures, luncheon Monday, Tuesday and Wednesday for members and guests, four group dinners and ten group luncheons, nineteen committee meetings, the annual reception and dinner-dance, an informal dance, a dinner-smoker and the anual meeting of the Woman's Auxiliary The attendance record for the past three annual meetings follows (figures are in the order 1936-1935-1934) total registration, 1658¬1609-1534, dinner dance, 666-577-543, dinner-smoker, 528-490-425, informal dance, 218-279-295, women's registration, 184-179-160, excursion, 183 (1936), luncheons, 1135-1132-1034 The Mineral Industry Education Division met at the Engineering Woman's Club Sunday afternoon and evening, Feb 16, with 55 present Three problems of current interest were discussed accrediting of engineering schools by the Engineers' Council for Professional Development, the question of what metallurgical engineers shall be taught, and the need for more attention to the industrial minerals in the curricula of mining schools A buffet supper was served In the evening the meeting became a joint session with the Industrial Minerals Division A second session of the Division was held on Wednesday afternoon, Feb 19, in the Engineering Societies Building The Sunday meeting was the only Annual Meeting technical session held outside the building

Jan 1, 1937

By C. S. Matthews, P. Hazebroek, F. Brons

A method has been developed for calculating the average pressure in a bounded reservoir. The reservoir is first divided into the individual drainage volumes of each well, by using the criterion that at steady state each individual drainage volume is proportional to a well's production rate. The average pressure in each drainage volume is then calculated by a method developed in the report. By volumetrically averaging these individual drainage volume pressures, the average pressure in the entire reservoir is obtained. To calculate the average pressure in each drainage volume, a correction is applied to the ordinary extrapolated pressure, i.e., the pressure obtained by extrapolating to infinite time the linear portion of the graph of closed-in pressure versus log[?t/(t + ?t)], where At is the closed-in time and t the production time. The correction, which is a function of the production time, is presented in graphical form for different shapes of the drainage area (horizontal cross section of the drainage volume). INTRODUCTION It is important to be able to find the volumetric average pressure in a reservoir so that the size of the reservoir may be determined from material balance calculations. It is also desirable to be able to find the approximate distribution of pressure within a reservoir for detection of fluid movement. 'The purpose of this paper is to present a method for calculating both the average reservoir pressure and the approximate distribution of pressure within a bounded reservoir — that is, a reservoir with no water drive. In reservoirs where the pressure builds up rapidly after wells are shut in, the determination of average pressure generally poses little problem, for one often need only average the final buildup pressures. It is when pressure buildup is slow that difficulties arise. For practical and economical reasons, the time allowable for closing in wells is limited. If at the maximum allowable closed-in time the pressure has not reached a con- stant value (and this is more often the case than is generally realized), calculation of average pressure presents difficulties. One method which has been used in the past for obtaining an "average" pressure involves the extrapolation to infinite closed-in time of curves of pressure as a function of closed-in time. Contour maps of these extrapolated pressures are used in arriving at an average pressure. In some cases wells are shut in for arbitrary Lengths of time, and contours of these arbitrary shut-in pressures are used for calculation. All these methods are somewhat arbitrary and suffer from lack of fundamental theoretical basis. For example, no precise meaning can be given either to the extrapolated or to the arbitrary shut-in pressure. Furthermore, there seems to be no reason why either set of pressures should be contourable. It was because of the vagueness of these old methods and their lack of a sound theoretical basis that, the present method for determining average reservoir pressure was developed. SYMBOLS A = drainage area of a well, cm' A, = drainage area of well j, cm2 A , = total drainage area of all wells in a reservoir, cm2 c = coefficient of compressibility of fluid at reservoir temperature, atm-1 f = hydrocarbin-filled porosity, fraction h = net thickness of formation, cm k = permeability of formation, darcy p = pressure at a well, atm p0 = initial reservoir pressure, atm p* = pressure obtained by straight-line extrapolation of linear portion of plot of p versus log [At/(t+ At)] p = volumetric average pressure inside drainage volume of a well q = volume flow rate at a well, cc/sec a: prevailing reservoir conditions q, = total volumetric flow rate from a reservoir, cc/sec at prevailing reservoir conditions rw = radius of well bore, cm T = kt/fµcA t = corrected time of production of a well, seconds, defined by

Jan 1, 1955

By Joseph Kelly

DEPLETION of anthracite resources in Lackawanna County, Pennsylvania, has forced the recovery of coal tracts formerly considered unminable. Chief among these are the large areas of outcrop coal lying on a heavy pitch on the mountain sides, abandoned thirty or forty years ago as too costly and too hazardous to mine. The old companies, in many instances, stopped their chambers 80 or 100 ft. hack from the actual outcrop, and then dropped back another 100 ft. before starting to rob the pillars, thus leaving large areas of coal. Recovery of this coal naturally leads to mining hazards not encountered years ago. The first step is to sink a slope from the surface with the pitch of the vein, and then drive counters across the pitch as close to the robbed area as proper protection for the lower side of the -counter will permit. Next, the old chambers are advanced or new ones driven in the solid, as the conditions require. As these chambers approach the outcrop the real safety work begins-that of preventing the miner from being buried or locked in by an avalanche of mud.

Jan 1, 1936

By G. Y. Chin, W. F. Hosford

It has long been recognized that, for a crystal to undergo an arbitrary shape change by crystallographic shear, at least five independent shear systems must be activated. Taylor1 analyzed the deformation of fcc crystals by {111}<110> slip and showed that the principle of minimum work could be used to determine which combination of slip systems might act to produce a given shape change. Later Bishop and Hi112 attacked the same problem by first identifying the state of stress which could physically activate five or more {111)(110) systems simultaneously and then invoking the principle of maximum virtual work to determine which combination would act for a given shape change. Taylor&apos;s general approach has been used to calculate the active slip systems and the relative strengths during axisymmetric flow of cubic crystals with other deformation modes, including {112}<111> slip3 and {111}<112> twinning.&apos; It is the purpose of this note to identify the stress states capable of activating simultaneously five or more of these systems. {111}<112> Twinning. The {111}<112> systems are defined in Table I and illustrated in Fig. 1. The sense of the shears are those for twinning in fcc crystals. The shear stress on any system may be written in terms of the stress components expressed relative to the 1,2,3 cube axes. For example, the stress For five or more of the systems ij to be activated simultaneously, it is necessary that Tij/T be equal to unity in at least five of the expressions [3], while Tij < T for the remainder. (These restrictions assume that the same shear stress T is required to activate each system and that the shear stress on any system cannot exceed that for its activation.) The discrete values of the stress components A through H which can satisfy these restrictions are listed in Table II together with the values of Tij/T for all systems. It may be noted that Tij/ T is not necessarily zero on all nonactive systems as found by Bishop and Hill for <110> {111} slip. In fact, for the case of twinning Tij/T may reach negative values as large as -2 for shear stress in the reverse direction, Table 11. This means that, if those stress states are to be valid, the shear stress to activate reverse twinning must be greater than twice that for forward twinning. Such stress states are probably valid in view of the directionality of the twinning process. There has been no documentation of reverse twinning thus far. It may be noted that the stress states fall into four classes, which may simultaneously activate eight, six, six, and five twin systems respectively. These solutions may be applied to the {112}(111)

Jan 1, 1970