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By H. A. Koch, R. L. Slobod

This paper discusses a new oil recovery process called the "miscible slug process." This process involves the injection of propane or LPG into the reservoir prior to gas injection. The operating conditions are maintained so that the slug is miscible with both the reservoir oil and the injected gas phase. Thus, a miscible phase displacement is achieved, and high recoveries are obtained from the area of the reservoir contacted. The most striking discovery of the laboratory study which included displacements from cores up to 123 ft in length was that relatively small slugs of LPG are effective over reservoir distances in the miscible displacement of oil. This makes possible the commercial application of the process. The important factors controlling the size of the slug are: (I) reservoir length, (2) reservoir fluid composition, and (3) reservoir pressure at the displacement front. Of lesser importance are the effects of injection rate and porous medium type. Questions which have not been completely answered deal with the effects of gravity and reservoir inhomogeneities on the process. INTRODUCTION One of the major problems facing the oil industry today is recovery of the large fraction of the discovered oil which will remain unrecovered unless some new processes are found. Water flooding has been one successful method used to increase recovery over natural depletion. But even successful water floods may leave 25 to 40 per cent of the pore space filled with residual oil. Recently a process which involves the use of thermal energy has been reported1,'*' which may prove useful in increasing recovery from shallow reservoirs containing highly viscous crudes. This paper deals with a new oil recovery process involving use of miscible phases to displace the oil from the reservoir. Miscible phase displacement oil has been an intriguing idea because the elimination of capillary effects in the reservoir leads to 100 per cent recovery in the areas contacted by the miscible displacing phase. The big deterrent in the past to the practice of miscible phase displacement has been the high cost involved in using large volumes of solvent in the process. The use of high pressure gas to achieve a miscible phase displacement of reservoir oil has been reported4,5 as one economically feasible method. However, this process calls for the maintenance of pressures above 3,000 psi at the dis-placement front. The process described in this paper is not restricted to high reservoir pressures; the only qualification being that the slug of material which is injected into the reservoir must be miscible with both the displaced oil 2nd the displacing phase. Two such cases are the use of an alcohol slug followed by water to achieve a mis-cable water flood, and the use of a low boiling hydrocarbon slug such as propane followed by gas to achieve a miscible gas drive. This paper will deal specifically with the propane dug process. This process will obtain a miscible phase displacement at pressures as low as about 1,200 psi at a reservoir temperature 01 150°F. The process is very simple. The propane displaces oil completely at one front because it is miscible with the oil, and gas displaces the propane completely at a second or trailing front because it is miscible with the propane. A small band of propane should thus remain wedged between the gas and oil phases. This idea has been called the "miscible slug process."6 Unfortunately, the propane slug will become diluted as it flows through the reservoir. This results from a mixing action since the fluids flow through a complex reservoir porous medium rather than simple capillaries. Fig. 1 diagrams a miscible slug displacement. The effects of some of

Jan 1, 1958

By J. T. Eash, N. B. Pilling

As a group, the alloys of iron, nickel and carbon are, in application, one of the most versatile of the ferrous alloy family, and while many investigations have been made of their properties and structure only a few attempts have been made to show the microstructural relation over a wide range of composition. It is neccssary at once to be clear about the kind of structural relation with which we are concerned. The equilibrium diagram provides quantitative data regarding structural make-up and its variation with temperature, but equilibrium structures often are far different from the technologically significant ones. This is particularly true of steels and cast irons, and some of the most important cornponents of their structure may have no place in the diagram of equilibrium. The fugitive phases that result from transFormation are dealt with in a highly detailed and descriptive manner in. the S-curves originated by Bain. The present effort lies in a still different field, which is to portray the relation between sweeping changes in composition and in resultant structure, under controlled but nonequilib-rium conditions of thermal treatment. The Guilletl diagram presented in 1903 was the initial attempt at such a correlation and frequent references to this diagram can be found in the literature even though later work has shown it to require considerable modification. Kase2 published in 1925 a diagram that was based upon observations of furnace-cooled melts: At low carbon contents this diagram agreed fairly well with experience in showing pearlite to exist in alloys containing up to 6 per cent nickel, and completely austenitic steels at 24 to 28 per cent nickel. At high carbon contents, however, irons were indicated as being martensitic at around 2 per cent nickel and austenitic at 12 per cent nickel; these interpretations are open to serious question. Based upon experience with steels and cast iron, Wickenden3 suggested a diagram that agreed fairly well with Kase's in the steel range but which presented a wide departure from it in the higher-carbon, high-nickel regions. In its broad picture, the Wickenden diagram is an approach to "practical" structures. Reference may be made at this point to the efforts of Marsh5 to construct an iron-nickel-carbon equilibrium diagram by the coordination of scattered data, which are described at length in the monograph cited. None of these preceding attempts has appeared to meet the need for an exposition of the structural sequence that is encountered in traversing the nickel-steel nickel-cast-iron range under what might be called ordinary departures from equilibrium. ExperimentaT. Conditions Since the intent has been to establish structures developed under arbitrary conditions, experimental details have perhaps more than usual importance. Three of these conditions should particularly be noticed: (I) the material was cast and not subjected

Jan 1, 1942

By J. T. Eash, N. B. Pilling

As a group, the alloys of iron, nickel and carbon are, in application, one of the most versatile of the ferrous alloy family, and while many investigations have been made of their properties and structure only a few attempts have been made to show the microstructural relation over a wide range of composition. It is neccssary at once to be clear about the kind of structural relation with which we are concerned. The equilibrium diagram provides quantitative data regarding structural make-up and its variation with temperature, but equilibrium structures often are far different from the technologically significant ones. This is particularly true of steels and cast irons, and some of the most important cornponents of their structure may have no place in the diagram of equilibrium. The fugitive phases that result from transFormation are dealt with in a highly detailed and descriptive manner in. the S-curves originated by Bain. The present effort lies in a still different field, which is to portray the relation between sweeping changes in composition and in resultant structure, under controlled but nonequilib-rium conditions of thermal treatment. The Guilletl diagram presented in 1903 was the initial attempt at such a correlation and frequent references to this diagram can be found in the literature even though later work has shown it to require considerable modification. Kase2 published in 1925 a diagram that was based upon observations of furnace-cooled melts: At low carbon contents this diagram agreed fairly well with experience in showing pearlite to exist in alloys containing up to 6 per cent nickel, and completely austenitic steels at 24 to 28 per cent nickel. At high carbon contents, however, irons were indicated as being martensitic at around 2 per cent nickel and austenitic at 12 per cent nickel; these interpretations are open to serious question. Based upon experience with steels and cast iron, Wickenden3 suggested a diagram that agreed fairly well with Kase's in the steel range but which presented a wide departure from it in the higher-carbon, high-nickel regions. In its broad picture, the Wickenden diagram is an approach to "practical" structures. Reference may be made at this point to the efforts of Marsh5 to construct an iron-nickel-carbon equilibrium diagram by the coordination of scattered data, which are described at length in the monograph cited. None of these preceding attempts has appeared to meet the need for an exposition of the structural sequence that is encountered in traversing the nickel-steel nickel-cast-iron range under what might be called ordinary departures from equilibrium. ExperimentaT. Conditions Since the intent has been to establish structures developed under arbitrary conditions, experimental details have perhaps more than usual importance. Three of these conditions should particularly be noticed: (I) the material was cast and not subjected

Jan 1, 1942

By D. W. Philips

The problem of strata control has of necessity been prominent in mining, and accounts of individual experience and research abound in the literature of most mining countries. Organized rescarch began in Britain by the formation of a committee, upon the recommendation of the Safety in Mines Research Board in 1923, to investigate methods of reducing accidents from falls of roof in coal mines. This committee examined current practice in supporting underground workings in the principal coal fields, and its findings have been published in the Board's publications.' - Toward the end of these investigations it was suggested by the Chief Inspector of Mines that committees of the several mining institutes be formed to examine the problem in their respective districts. The formation of these committees marked a new phase in research on strata control. The work, though being done in the separate coal fields, followed a scheme outlined by the chief Mining Engineer of the Research Board, and was coordinated by his office. FACTORS INVOLVED The subject is being considered in Britain from two aspects: 1. To provide. by supports, means to prevent falls of fractured ground. 2. TO develop methods whereby the strata can be maintained in the least broken condition and thus make falls less likely. The first aspect is the one that will most likely lead to immediate results from the accident standpoint. It depends largely upon the assimilation of already acquired knowledge regarding the properties of supports and their correct application. The second aspect is perhaps more fundamental, and involves also the properties of the strata and their behavior under the influence of the forces induced by mining. The problem involves many natural factors that cannot be controlled, such as depth, inclination and thickness of seam and nature of the rocks. Other factors, which may be considered to a certain extent controllable, cannot be varied readily and with sufficient frequency to facilitate easy investigation; some of these are: rate of advance of the working face, length of face, direction of the face with regard to cleat and inclination, system and arrangements of rock packing behind longwall faces. In any given case the very many factors present have a mutual influence upon the effects produced in the mine. For example, a study of the effects of rate of advance cannot be done without considering the effects of other factors such as nature of strata, direction of face and quality and arrangement of supports. Conversely, the length and direction of a working face

Jan 1, 1947

By Charles B. Dudley

THEO. N. ELY, Esq., Superintendent of Motive Power. DEAR SIR: It is now nearly three years since my first report to you on the subject of steel rails was written. That report, as you will remember, dealt principally with the question of the relation between the chemical composition and physical properties of steel rails and their power to resist crushing and fracture in actual service. Other matters were referred to or touched upon in that report, but the main question was, Why do some rails crush or break in service while others do not? You will doubtless remember that the principal conclusion arrived at was, that the softer rails are less liable to crush or break in service than the hard ones. Or, in other words, so far as conclusions could be drawn from the chemical analysis and physical test of 25 samples of steel rails which had actually been in service, these conclusions were that those rails which have the smaller amounts of carbon, phosphorus, silicon, and manganese are less liable to crush or break in service than those which have larger amounts of these elements. Or, again, looked at in the light * At the Lake George meeting of the Institute in October, 1878, I had the honor of presenting to the Institute, through the kind permission of the officers of. the Pennsylvania Railroad Company, the results of a study of twenty-five samples of steel rails, which had all been in actual service. Considerable discussion followed the publication of that paper, and there seemed to be a strong disinclination, especially on the part of the steel-rail manufacturers, to accept the conclusions presented. One of the principal objections urged against the conclusions drawn was, that they were based on too few samples; in other words, that no conclusions safe to act upon could be drawn from the examination of twenty-five rails. In view of this criticism it was decided to repeat the investigation with a larger number of samples, and with the aid of the experience gained in the first investigation. The results of this second study of steel rails are, by the permission of the officers of 'the Pennsylvania. Railroad Company, herewith presented to the Institute, with the sincere desire that they may aid in adding to our knowledge of this most important product. Like the previous paper, this is in the form of a report to one of the officers of the company, which will account for the style, and the manner of presenting the data

Jan 1, 1881

By R. F. Mehl, C. S. Barrett, A. H. Geisler

THE mechanism by which the super-saturated solid solution of 20 per cent silver in aluminum decomposes has been reported in two previous publications. 1,2 The analysis of streaks in Laue photograms showed that thin platelike particles (the so-called Guinier-Preston aggregates) form upon {111} matrix planes in the early state of decomposition.1 The size of these particles is larger, the higher the temperature at which they form in accordance with nucleation theory;3 at all temperatures they increase in size with aging.2 When they have grown thick enough to afford three-dimensional diffraction, powder X-ray photograms show that the structure is not that of the equilibrium precipitate y but that of a transition lattice ?'. Since this is also platelike in shape and also lies upon the {111} matrix planes with the {110} direction in the matrix parallel to the {11.0} direction in ?', it seems likely that the original platelet particles are merely ?' in an early stage of formation, and thus finely dispersed. The atomic spacing and pattern on the {00.0} plane in ?' are identical with those on the {111} matrix plane to which it lies parallel, and since comparable directions in the two planes are parallel, the lattices are thus coherent at the interface; an analogous coherence has been found in aluminum-copper alloys.3 The structures ?' and y are both close-packed hexagonal and are oriented alike; they differ only in interatomic distances. Consequently, ? ' is regarded as y strained from its ordinary dimensions by the lattice coherency between it and the matrix.2 The lattice movements that transform the matrix to the equilibrium precipitate appear to be very simple.2 The mechanism of precipitation in this system is simpler and is more precisely known than that in any other system,2,4 hence the system should be a profitable one for the difficult task of correlating property changes with structural alterations. The present report provides additional studies on powder X-ray diffraction from alloys of 10 per cent, 20 per cent and 30 per cent Ag, and presents the correlation of crystallographic changes with changes in microstructure, Brinell hardness and electrical resistivity of the three alloys. PREPARATION OF SAMPLES AND EXPERIMENTAL PROCEDURE The preparation of the 20.2 per cent Ag alloy has been described.2 Alloys containing 9.9 and 29.7 per cent Ag were prepared from 99.97 per cent Al and 99.99+ per cent Ag using a similar procedure. Rolled

Jan 1, 1943

The first record of coal anywhere in the Appalachian regions of which we now know is along the north fork of the Potomac River, above the mouth of Savage River, on a map entitled, A Plan of the upper Part of Potomack River call Cohongorooto Survey'd in the Year 1736. Benj. Winslow1 two "Cole-mines" are shown, one just above Green Island and the other just below Hopwood Run. Winslow was in charge of a surveying party to "Lay out the Bounds of the Northern Neck of Virginia," or Lord Fairfax's Grant, his party locating the liver above the mouth of the Shenandoah, while another party located it below that point, and then around Chesapeake Bay and up the Rappahannock River. The results of the two surveys were combined in a map called, The Courses of the Rivers Rappahannock and Potowmack, in Virginia, as surveyed according to Order in the Years 1736 & 1737, made by Wm. Mayo, which does not show the "cole-mines;" this map was used with the report of the boundary commission which was sent to London in the case before the Privy Council. The boundary line decided upon was surveyed on the ground in 1746, just ten years after the original one. The results of this work were shown on A Map of the Northern Neck in Virginia, According to an Actual Survey begun in the Year MDCCXXXVI, and ended in the Year MDCCXLVI, Drawn by Peter Jefferson And Robert Brooke, Surveyors, which is now in the Colonial Office in London. This map shows the "cole-mines" in the same location as on Winslow's map, and taken from his record, of course, as that part of the survey was not retraced. Faulkner, who examined all of these records in his review of the case in 1832, refers to the notes of the original survey as "No. 10. The original field notes of the survey of the Potomac River, and the mouth of the Shenandoah to the head spring of said Poto¬mac River, by Mr. Benjamin Winslow." While the notes of the 1746 survey are in the London records and photographic copies of them are in several libraries in this county, the notes of the 1736 survey cannot be found anywhere. (For Chas. J. Faulkner's report, dated Nov. 6, 1832, see Kercheval's History of the Valley of Virginia, pp. 160-173. Jas. W. Foster's Maps of the First Survey of the Potomac River, 1736-37, in Wm. and Mary College Quarterly History Mag., April, 1938, gives a complete history

Jan 1, 1942

By D. E. Hartman, J. M. Roberts

A detailed study of the temperature dependence of the critical stress, TB . necessary to cause a damping loss of 1.41 x 10-6 g mm per mm3 in pre-strained magnesium single crystals has been carried out in the temperature range 82" to 320°K. Tb is found to be linearly related to the reciprocal of the ahsol~cte temperature in this temperature range. The temperature dependence of the anelastic limit, TA, defined as the stress necessary to form open unidirectional stress-cycle damping loops has been measured between 82" to 320°K. ta is found to be independent of temperature in this range. It is suggested that TA is determined by the stress necessary to activate Frank-Read dislocation NUMEROUS studies1-9 have been reported concerning the details of yielding and microstrain near and below the generally discussed macroscopic flow stress. The work of Kramer and coworkers10-13 is helpful in determining certain qualitative facts. concerning preyield phenomena. The absence of a stress-dependent delay time in annealed copper and aluminum single crystals between 83" and 298 suggests that the preyield mechanism is temperature-independent in this range. This is not sources. The temperature and stress dependence of the effective activation volume (Veff) has been measured in the damping-loop region in prestrained magnesium single crystals. The data suggests that Veff/kT is approximately constant, equal to 2 4.5 sq mm per g in the temperature range 130°to 298°K, and that Veff is almost independent of stress at stress levels greater than 10 g per sq mm. A restricted model is developed to predict the stress (rb) necessary to cause a small fixed damping loss as a function of temperature, effective activation volume, and dislocation density. Various possible dislocation mechanisms controlling TB are discussed. true for zinc, p brass, and iron crystals in the same temperature range, however, since the existence of a stress-dependent delay time suggests a thermally activated preyield mechanism in these materials.10-13 The work by Vreeland et a1.14 confirms the work by Liu et al.ll on delayed yielding in zinc single crystals. A stress-dependent delay time between 82" and 298°K was found12 in prestrained aluminum and copper crystals. These results suggest that the preyield phenomena may be different between prestrained and unstrained or annealed crystals. The temperature dependence of the yield point at 2 x 10-8 plastic strain found by Rosenfield and Averbach3 in annealed copper and aluminum is not in agreement with the delay-time data, unless possibly the delayed yielding experiment only de-

Jan 1, 1964

By D. W. Philips

The problem of strata control has of necessity been prominent in mining, and accounts of individual experience and research abound in the literature of most mining countries. Organized rescarch began in Britain by the formation of a committee, upon the recommendation of the Safety in Mines Research Board in 1923, to investigate methods of reducing accidents from falls of roof in coal mines. This committee examined current practice in supporting underground workings in the principal coal fields, and its findings have been published in the Board's publications.' - Toward the end of these investigations it was suggested by the Chief Inspector of Mines that committees of the several mining institutes be formed to examine the problem in their respective districts. The formation of these committees marked a new phase in research on strata control. The work, though being done in the separate coal fields, followed a scheme outlined by the chief Mining Engineer of the Research Board, and was coordinated by his office. FACTORS INVOLVED The subject is being considered in Britain from two aspects: 1. To provide. by supports, means to prevent falls of fractured ground. 2. TO develop methods whereby the strata can be maintained in the least broken condition and thus make falls less likely. The first aspect is the one that will most likely lead to immediate results from the accident standpoint. It depends largely upon the assimilation of already acquired knowledge regarding the properties of supports and their correct application. The second aspect is perhaps more fundamental, and involves also the properties of the strata and their behavior under the influence of the forces induced by mining. The problem involves many natural factors that cannot be controlled, such as depth, inclination and thickness of seam and nature of the rocks. Other factors, which may be considered to a certain extent controllable, cannot be varied readily and with sufficient frequency to facilitate easy investigation; some of these are: rate of advance of the working face, length of face, direction of the face with regard to cleat and inclination, system and arrangements of rock packing behind longwall faces. In any given case the very many factors present have a mutual influence upon the effects produced in the mine. For example, a study of the effects of rate of advance cannot be done without considering the effects of other factors such as nature of strata, direction of face and quality and arrangement of supports. Conversely, the length and direction of a working face

Jan 1, 1947

By Strathmore R. B. Cooke, Paul B. Queneau

The kinetics of the dissolution of scheelite in basic solutions of carbonate and fluoride were investigated. Work was also done with solutions of alkali metal chloride, hydroxide, phosphate, and mixtures of phosphate and fluoride. The study was conducted at 90 psig and at temperatures ranging from 100" to 155"C, and both natural and synthetic scheelite single crystals were used. The interaction of scheelite with carbonate solutions generally followed a parabolic rate law, and resulted in the formation of calcite crystals on the attached surface. The transport of the CO3 through the calcite to the scheelite surface was probably rate -controlling between 115" and 135°C. Positive deviations from Parabolic behavior were observed at 100" and 155?C, presumably due to the calcite grain growth, and to the formation of a noncoherent calcite layer, respectively. The leaching of scheelite with alkaline fluoride solution produced a calcium fluoride reaction-product which covered the surface as a sheet-like layer. Below 110°C the dissolution rate was parabolic; at 115" and 135°C it was linear. In the linear region, the periphery of the calcium fluoride layer lifted from the sample. Only the lower layers, braced by the scheelite substratum, remained unbroken. This bound layer, approximately constant in thickness, limited the rate at which the fluoride could attack the scheelite. The phosphate and phosphate-fluoride leaches dissolved scheelite at a linear rate; solid reaction products were respectively hydroxyapatite and fluorapatite. Sodium hydroxide and chloride were not effective leachants, and did not produce observable solid reaction products. THE feasibility of pressure leaching of scheelite pulps with basic solutions containing either carbonate or fluoride ions, as indicated by the following equations, has been established by previous investigators:1-3 200°C CaWO4s + CO37q - CaCO3s+WO4aq [1] 200°C CaWO4s + 2Faq ^ CaF2s+WO4aq [2] It is to be noted that a commercial process employing reaction [I] is in use at Union Carbide's operation near Bishop, Calif. In the earlier work referred to, it was apparent that the precipitation of the solid reaction-product on the scheelite particles sharply reduced contact of the leachant with the scheelite. We have extended this work by studying the kinetics of dissolution employing single crystals of scheelite of known exposed surface area. In addition, the effectiveness of solutions of phosphate, of phosphate with fluoride, and of chloride and hydroxide as leachants for scheelite have been investigated. EXPERIMENTAL Apparatus. The low-pressure autoclave used in this wo (100 psig maximum) is essentially that of J. Halpern4 as modified by F. A. Forward.5 Constructed entirely of 316 stainless steel, its maximum capacity is 4000 ml. The temperature in the autoclave was regulated to *0.5?C by a Wheelco thermocouple controller. A 4-in. diam, three-bladed impeller, circulated solution upward against the sample, and a strobo-scope permitted holding the stirring rate to *15 rpm. A check valve and a "pop" type safety valve were installed, a sample holder was fixed to an externally adjustable rod, and a water-cooled condenser was placed in series with the sampling valve. The stainless steel sample holder, Fig. 1, held the epoxy-mounted sample (f) five mm above the impeller. Tightening the tapered sleeve (d) positioned the epoxy-resin specimen-mount by forcing it back against the spacer (c). The mount (e) was 1 in. diam. Sample Description and Preparation. Both synthetic and natural scheelite crystals were leached. The former were single crystals of 99.95 pct purity which were leached approximately perpendicular to the "c" axis; they were supplied by the Gallard-Schlesinger Chemical Manufacturing Corp. The natural crystals, supplied by Minerals Unlimited, were transparent, analyzed 99.8 pct CaWO4, and in ultraviolet illumination exhibited the characteristic blue glow of substantially molybdenum-free scheelite. To permit exposure of an essentially constant surface area of scheelite to the leachant, the samples were mounted in Scotchcast Brand Resin No. 250 (Minnesota Mining and Manufacturing Co.). The mount-

Jan 1, 1970

By Raymond D. Sloan

Dropping from third position among the oil-producing states of the nation in 1942, Oklahoma ranked fourth in 1943 with a total output of 121,431,ooo bbl., a decline of 11.9 per cent from the previous year's total of 137,792,000 bbl. Production ranged from a maximum of 346,900 bbl. daily in February to a low of 322,800 bbl. daily in October, and at the close of the year was averaging 330,000 bbl. daily from 52,096 wells, or an average of 6.3 bbl. per well. The state's well, average at the close of 1942 amounted to 6.7 bbl. At the turn of the year, it appeared reasonable to assume that Oklahoma's producing rate would continue its downward trend during 1944 unless reversed by greater than normal discoveries. To Jan. I, 1944, Oklahoma has produced 5,188,466,-ooo bbl. of crude oil. Development and Exploration A total of 1287 wells was drilled in 1943, representing an increase of 77 wells, or 6.4 per cent over the preceding year's total of 1210. Of the wells completed, 587 (45.6 per cent) were oil wells with an average initial production of 319 bbl.; 127 (9.9 per cent) were gas wells with an average initial of 8,639,000 cu. ft.; and 573 (44.5 per cent) were dry. A total of 333 wildcats was drilled (25.9 per cent of all wells), an increase of 13 (4.1 per cent) over the 1942 total of 320. Of the*wildcats drilled, 46 (13.8 per cent) were oil wells with an average initial production of 235 bbl.; 27 (8.1 per cent) were gas wells with an average initial of 8,249,-ooo cu. ft.; and 260 (78.1 per cent) were recorded as failures. Reconditioning of wells throughout the state led to a total recompletion of 230 wells. Of these, 131 were oil wells with a combined initial production of 8360 bbl., 31 were gas wells with a total initial of 211,950,000 cu. ft., and 68 were abandoned. The 46 successful wildcats opened 33 new producing areas, one of which produced from three separate horizons, and led to the development of production in new horizons in 11 old pools. While none of the 1943 discoveries have as yet developed into what might be termed major pools, the West Moore pool, in Cleveland County, and the West Edmond pool, in Oklahoma County, are the two outstanding new areas opened to production. The West Moore pool, sec. 29, 10 N., 3 W., discovered in December 1943, was completed for an initial production of 248 bbl. of oil in 3 hr. through small tubing chokes in the second Wilcox sand from 8787 to 8800 ft. One well was drilling in this area at the close of the year. The West Edmond pool, an April discovery in sec. 32, 14 N 4 W., is producing from the Hunton lime at a depth of approximately 6950 ft., the world's deepest Hunton lime pool. To the end of the year, 15 oil wells had been completed, with no gas wells and no failures. Accumulated production to Jan. I, 1944, amounted to 474,300 bbl. of 41" gravity oil. There were 29 active operations in the field at the end of the year. The West Edmond and West Moore pool discoveries established the first commercial production west of the Granite Ridge,

Jan 1, 1944

By Raymond D. Sloan

Dropping from third position among the oil-producing states of the nation in 1942, Oklahoma ranked fourth in 1943 with a total output of 121,431,ooo bbl., a decline of 11.9 per cent from the previous year's total of 137,792,000 bbl. Production ranged from a maximum of 346,900 bbl. daily in February to a low of 322,800 bbl. daily in October, and at the close of the year was averaging 330,000 bbl. daily from 52,096 wells, or an average of 6.3 bbl. per well. The state's well, average at the close of 1942 amounted to 6.7 bbl. At the turn of the year, it appeared reasonable to assume that Oklahoma's producing rate would continue its downward trend during 1944 unless reversed by greater than normal discoveries. To Jan. I, 1944, Oklahoma has produced 5,188,466,-ooo bbl. of crude oil. Development and Exploration A total of 1287 wells was drilled in 1943, representing an increase of 77 wells, or 6.4 per cent over the preceding year's total of 1210. Of the wells completed, 587 (45.6 per cent) were oil wells with an average initial production of 319 bbl.; 127 (9.9 per cent) were gas wells with an average initial of 8,639,000 cu. ft.; and 573 (44.5 per cent) were dry. A total of 333 wildcats was drilled (25.9 per cent of all wells), an increase of 13 (4.1 per cent) over the 1942 total of 320. Of the*wildcats drilled, 46 (13.8 per cent) were oil wells with an average initial production of 235 bbl.; 27 (8.1 per cent) were gas wells with an average initial of 8,249,-ooo cu. ft.; and 260 (78.1 per cent) were recorded as failures. Reconditioning of wells throughout the state led to a total recompletion of 230 wells. Of these, 131 were oil wells with a combined initial production of 8360 bbl., 31 were gas wells with a total initial of 211,950,000 cu. ft., and 68 were abandoned. The 46 successful wildcats opened 33 new producing areas, one of which produced from three separate horizons, and led to the development of production in new horizons in 11 old pools. While none of the 1943 discoveries have as yet developed into what might be termed major pools, the West Moore pool, in Cleveland County, and the West Edmond pool, in Oklahoma County, are the two outstanding new areas opened to production. The West Moore pool, sec. 29, 10 N., 3 W., discovered in December 1943, was completed for an initial production of 248 bbl. of oil in 3 hr. through small tubing chokes in the second Wilcox sand from 8787 to 8800 ft. One well was drilling in this area at the close of the year. The West Edmond pool, an April discovery in sec. 32, 14 N 4 W., is producing from the Hunton lime at a depth of approximately 6950 ft., the world's deepest Hunton lime pool. To the end of the year, 15 oil wells had been completed, with no gas wells and no failures. Accumulated production to Jan. I, 1944, amounted to 474,300 bbl. of 41" gravity oil. There were 29 active operations in the field at the end of the year. The West Edmond and West Moore pool discoveries established the first commercial production west of the Granite Ridge,

Jan 1, 1944

By A. W. Ruff

In the last ten years large gains have been made in the field of blasting. These gains have been both in the theoretical and in the practical application of explosives. One of the most publicized changes in the practical application of commercial explosives is the now-prevalent use of ammonium nitrate and fuel oil. The new water-compatible mixture of ammonium nitrate and TNT should also have a large field of application. Most of the theoretical gains have been concerned with the actual rock breakage obtained with explosives. In the last ten years, large gains have been made in the field of blasting. These gains have been both in the theoretical and in the practical application of explosives. One of the most publicized changes in the practical application of commercial explosives is the now prevalent use of ammonium nitrate and fuel oil. The new water-compatible mixture of ammonium nitrate and TNT should also have a large field of application. Most of the theoretical gains have been concerned with the actual rock breakage obtained with explosives. EXPANDING-GAS-BUBBLE THEORY There has been an outstanding change in the conception of how an explosive produces rock failure. The early theory which explained rock breakage was correct for some of the first force-producing agents such as quicklime. When a drillhole was filled with quicklime and water added, the material expanded and steam was generated. The expanding material and the steam exerted a slowly increasing pressure on the sides of the drillhole. The force broke the rock by splitting. Because a great quantity of gas was produced by the deflagration of gunpowder, the same process of the gas pressure splitting the rock was assumed to take place. Even after 80 years of use, the rock-breaking action of all high explosives was thought to be caused entirely by the expanding gas bubble which tore and split the rock. Fig. 1 shows a vertical section through an exploding spherical charge of high explosive. It illustrates the early theory which attributed the entire breakage to the expanding gas bubble. The basic idea of this theory is that the breakage started at the explosive charge and progressed toward the free face. REFLECTED-WAVE THEORY me expanding-gas- bubble theory has gradually been replaced by the reflected-wave theory. The reflected-wave theory explains rock failure by explosives in the following manner: The enormously high gas pressure in the drillhole produces fracturing of the rock on the sides of the drillhole. The fracturing around a 1 1/2-in. diam drillhole is usually confined to an area with about a 6-in. diam. In the fractured area, the pressure of the gas is reduced until the force on the walls is no longer capable of fracturing the rock. Although the gas pressure does not continue to fracture the rock, the pressure imposes a high impact stress on the wall rock. In addition, a pressure or seismic pulse is produced in the rock. The pulse is probably entirely a compression pulse when originally produced in the rock. But, with distance, part of the energy of the compression pulse is transformed into a tension pulse which follows the compression pulse. The explosive charges and burdens used in conventional mine blasting are such that the tension pulse of the original wave is probably too small to be considered in a practical evaluation of the theory. The compression pulse is transmitted through the rock with no associated fracturing.

Jan 1, 1961

By E. Hein, R. A. Dodd

The fatigue softening of prior strain-hardened poly crystalline copper has been determined by measuring changes inflow stress resulting from fatigue treatments. Tensile fatigue does not soften the metal, while the softening due to reversed stress fatigue depends on the extent to which fully reversed stressing is approached. True tensile fatigue is shown to be possible only in the case of long wire specimens. Annealing following fatiguing of strain-hardened metal shows that tensile fatigue is very effective in modifying normal recovery and recrys-tallization. OBSERVATIONS by Ludwik and Scheu,' Kenyon,' Polakowski and Palchoudhuri, Kemsley, and Broom and Ham,596 have shown beyond doubt that strain-hardened metals are softened by reversed-stress fatigue. To some extent the softening might be associated with a rearrangement of dislocations, but the results of Broom and Ham5 on the temperature dependency of the fatigue hardening of annealed copper, and those of McCammon and Rosenberg7 on the partial recovery at 100'C of fatigue-hardened copper suggest that point defects may be involved. Since lattice vacancies are the only species of point defect present in appreciable quantities in thermody-namic equilibrium, most attention has been accorded them in the various theories advanced to explain hardening and softening effects resulting from fatigue. Precise mechanisms remain uncertain, and while some effects may be due to single vacancies, defects arising from vacancy clusters may also contribute to the changes in properties. All types of plastic deformation result in point-defect formation, one frequently invoked formation mechanism involving the nonconservative motion of jogs formed by the intersection of, for example, mixed dislocations. But since fatigue deformation is thought to be a particularly effective way of forming point defects, additional mechanisms peculiar to fatigue must be sought. One possibility is that jogged loops contract during the unloading cycle to give rows of vacancies. A local high-vacancy concentration could conceivably promote polygonization and recovery by climb, and, therefore, could explain the fatigue softening of strain-hardened metals under appropriate conditions of temperature. Conversely, the experiments by Broom and Ham6 on the fatigue hardening of copper single crystals involving subsequent tensile testing of previously fatigued specimens, resulted in the observation of distinctly lowered yield points. This could be due to vacancy atmospheres associated with dislocations, possibly augmented by jogs on dislocations. In addition to the yield-point phenomenon, a post-yield hardening was observed, with a temperature-dependence resembling the friction-hardening of irradiated metals,9 this again suggesting a point-defect mechanism. An important development in this field has been the observation,10 by transmission electron microscopy, of high concentrations of prismatic loops in fatigued aluminum. By analogy with quenched aluminum filmso11 it seems certain that these loops originate in the collapse of vacancy clusters formed during deformation. Metals, such as copper, having a relatively low-stacking fault energy would tend to produce Frank sessile loops (in practice tetrahedral defects with stacking faults on all (111) planes) rather than the glissile prismatic loops observed in aluminum. The manner in which these various defects affect the different aspects of fatigue behavior has not yet been fully investigated. The present work was designed principally to indicate whether tensile fatigue gives rise to hardening and softening effects similar to those associated with reversed stress fatigue, and with the hope of providing additional information on the contribution of point defects to fatigue deformation. Despite the uncertain ty often associated with the interpretation of resistivity data, i.e., the relative contribution of stacking faults and other defects, such measurements are conveniently employed to study point defect pheno-

Jan 1, 1962

By M. King Hubbert, David G. Willis

A theoretical examination of the fracturing of rocks by means of pressure applied in boreholes leads to the conclusion that, regardless of whether the fracturing fluid be of the penetrating or non-penetrating type, the fractures produced should be approximately perpendicular to the axis of least stress. The general state of stress underground is that in which the three principal stresses are unequal. For tectonically relaxed areas characterized by normal faulting, the least stress should be horizontal; the fractures produced should be vertical with the injection pressure less than that of the overburden. In areas of active tectonic compression, the least stress should be vertical and equal to the pressure of the overburden; the fractures should be horizontal with injection pressures equal to or greater than the pressure of the overburden. Horizontal fractures cannot be produced by hydraulic pressures less than the total pressure of the overburden. These conclusions are compatible with field experience in fracturing and with the results of laboratory experimentation. INTRODUCTION The hydraulic-fracturing technique of well stimulation is one of the major developments in petroleum engineering of the last decade. The technique was introduced to the petroleum industry in a paper by J. B. Clark,' of the Stanolind Oil and Gas Co. in 1948, and since then its use has progressively expanded so that by the end of 1955 more than 100,000 individual treatments had been performed. The technique itself is mechanically related to three other phenomena concerning which ah extensive literature had previously developed. These are: (1) pressure parting in water injection wells in secondary-recovery operations, (2) lost circulation during drilling, and (3) the breakdown of formations during squeeze-cementing operations, all of which appear to involve the formation of open fractures by pressure applied in a wellbore. The most popular interpretation of this mechanism has been that the pressure had parted the formation along a bedding plane and lifted the overburden, notwithstanding the fact that in the great majority of cases where pressures were known they were significantly less than those due to the total weight of the overburden as determined from its density. Prior to 1948, this prevalent opinion had already been queried by Dickey and Andresen,2 in a study of pressure parting; and by Walker,"' who, in studies of squeeze cementing, pointed out that the pressures required were mostly less than those of the overburden, and inferred that the fractures should be vertical. J. B. Clark,' in his paper introducing hydraulic fracturing, and later Howard and Fast," and Scott, Bearden, and Howard," also of Stanolind, postulated that the entire weight of the overburden need not be lifted in producing horizontal fractures, but that it was only necessary to- lift an "effective overburden," requiring a correspondingly lower pressure. Hubbert,7 in discussing the paper by Scott and associates, pointed out that the normal state of stress underground is one of unequal principal stresses; and in tectonically relaxed areas, characterized by normal faults, the least stress should be horizontal. Therefore, in most cases, fracturing should be possible with pressure less than that of the overburden and, moreover, such fractures should be vertical. Harrison, Kieschnick, and McGuire,8 also on the expectation that the least principal stress should be horizontal, argued strongly in favor of vertical fracturing. Scott, Bearden, and Howard6 observed that, when

Jan 1, 1958

By V. A. Phillips, M. Safdar Ali

The tensile properties of Cu-0.20 pct Al, Cu-0.45 pct Mg, Cu-0.27 pct Cr, and Cu-0.22 pct Be solid-solution alloys were studied at -196°, 18°, 2509 and 500°C on wires internally oxidized at 900°and 1000°C. Internal oxidation produced marked increases in yield strength relative to pure copper, particularly at low temperature. The yield strength decreased with increase in temperature to a much greater extent than predicted by current theories of dispersion hardening. The strengthening effect of internal oxidation persisted throughout the entire stress-strain curve, but since a marked loss of ductility occurred at all test temperatures, only modest increases in tensile strength were realized. INTERNAL oxidation involves the preferential oxidation of a baser solute in a relatively noble solvent by diffusion of oxygen into the solid. Since the process is diffusion controlled, it is expedient to carry it out at a high temperature often approaching the melting point. The amount of solute is limited and the oxidation conditions adjusted so that oxygen diffuses inwards to meet the solute causing reaction to progress inwards and form subscale instead of only sur face scale. The reaction can be made to occur throughout the thickness of thin material producing a fine dispersion of oxide inside the metal. C. S. Smith1"3 appears to have been the first to recognize the true nature of the process of internal oxidation. Although extensive kinetic and metal-lographic studies followed by Rhines and coworkers,4-5 Chaston7 andMeijering and Druyvesteyn8 working independently were the first to report hardening effects due to internal oxidation. These results prompted detailed studies of the effects of internal oxidation on the hardness, tensile, creep, and fatigue properties of a number of alloys by G. C. Smith and coworkers9-12 in England, on hardness by Gottardi13 in Italy, and on hardness and tensile properties by wood14 in the U.S.A. Publications dealing with other aspects of internal oxidation will not be reviewed here. Internal oxidation of solutes having a high affinity for oxygen provides a means of obtaining very fine dispersions stable in a suitable atmosphere at temperatures well above those at which most age harden ing precipitates coalesce. The process therefore offers a method of producing thin sheet and wire of possibly outstanding creep resistance. The size limitation might be overcome by internally oxidizing porous powder compacts made from alloy powder before final sintering. PREPARATION OF ALLOYS The alloy ingots, which were approximately 5 1/2 in. long by 1 in. diam, were made up in a high-frequency furnace, using as raw materials OFHC copper, 99.99 + pct Al, high-purity magnesium, and master alloys of Cu-4.19 pct Be and Cu-10 pct Cr. The alloys containing aluminum and magnesium were melted ii closed high-purity graphite crucibles and solidified in the crucible which was lowered partly Out of the coil to give a hot-top effect. The alloys containing beryllium and chromium were made by melting in vacuo in alumina-lined sillimanite crucibles. The magnesium, chromium, and beryllium

Jan 1, 1960

By R. T. Hukki

This paper presents a series of equations for mechanical ball wear, relating parameters of ball size, mill speed, and mill diameter. The fundamental equation, Eq. 12, presented here is introduced to correlate these basic parameters and thus define and clarify the concept of ball wear. This equation is offered as a general rule, which may be modified to apply to individual problems of grinding. BALL wear as observed in grinding installations is the combined result of mechanical wear and corrosion. Corrosion should be a linear function of the ball surface available. Ball corrosion, however, has been studied so little that its effect, although of great importance, cannot be included in the analyses given here. In a separate paper' it is shown that 1 n = 0.7663 np----=— rpm [1] vD P = c, np D kw [2] T=c²(np)n De tph [3] In these equations n — actual mill speed, rpm np = calculated percentage critical speed D = ID of mill in feet P = power required to operate a mill, kw T = capacity of a mill, tph C¹ and c² - appropriate constants in = exponent of numerical value of 1 5 m 1.5 Exponent m is the slope of a straight line on logarithmic paper relating mill speed (on the abscissa) and mill capacity (on the ordinate). It is generally accepted, although not sharply defined, that ball wear in mills running at low (cascading) speeds is a function of the ball surface available. Accordingly, the wear of a single ball may be considered to be a homogeneous, linear function of its surface and of the distance traveled. Thus dw = f¹(d2) . f2(ds) [4] where dw is the wear of a single ball in time dt, d the diameter of the average ball in ball charge, and ds the distance traveled by the ball in time dt. Indicating that ds - a D n dt, the wear of the average ball in time dt becomes dw = f¹(d2) . f2(Dn dt) 1 --- f¹ (d1) f² (D c3 np-----— dt) \/D = c,d² n, D dt The rate of wear of the average ball is given by dw/dt. dw/dt = c, d² np D lb per hr [5] The weight of the ball charge per unit of mill length is a function of D The number of balls of size d in the ball charge is = f³(D2)/f4(d³). The rate of wear of the total ball charge equals the number of balls times rate of wear of the average ball. Thus rate of total ball wear = — . (dw/dt) w. c, . (l/d) . n,, D lb per hr [6] which is the equation of ball wear in low speed mills. In a mill running at a low speed, grinding is the result of rubbing action within the ball mass and between the ball mass and mill liners. When the speed of the mill is gradually increased toward the critical, the impacting effect of freely falling balls becomes increasingly prominent in comparison with the rubbing action. Reduction of ore takes place partly by rubbing, partly by impact. The share of the freely falling balls in the reduction of ore reaches its practical maximum at a speed somewhat less than the critical; at that speed grinding by rubbing has decreased to a low value. It may be reasonable to think that size reduction by freely falling balls should reach its theoretical maximum at the critical speed, if the fall of the balls were not hindered by the shell of the mill beyond the top point; grinding by rubbing would cease at the critical speed. As a first approximation, wear of freely falling balls may be considered to be a homogeneous, linear function of the force at which they strike pieces of rock and other balls at the toe of the ball charge. The force equals mass times acceleration. The mass of a ball is a function of d3 and its acceleration is a function of the peripheral speed of the mill. The wear of a single ball of size d representing the average ball in a ball charge will therefore be w¹ = f3(F) = f (d3) f7 (v). [7] Indicating that v = D n, and n = c³ np 1/vD, Eq. 7 becomes W1 = cn d3 np Do.5 lb per hr. [8] Total wear of the ball charge equals number of balls times the wear of the average ball. Number of

Jan 1, 1955

By W. K. Hartley

INTRODUCTION AND HISTORY The Kimberley Division of De Beers Consolidated Mines Limited includes four underground mines in Kimberley viz. Dutoitspan, Bultfontein, De Beers and Wesselton Mine and two outside mines, viz. Finsch, an opencast operation and Koffie- fontein, an underground operation. Figure 1 is a map of the Kimberley district showing the location of these mines. Finsch is 150km west of Kimberley and Koffiefontein is 80km south of Kimberley. The head office of the De Beers Company is in Kimberley. The diamond sorting and trading companies for diamonds produced in South Africa are also in Kimberley. Diamonds were first discovered in the Kimberley area at Hopetown in 1866 on the banks of the Orange River. This, and later finds, triggered the start of the great "Diamond Rush". In 1870 the Jagersfontein (now closed) and Dutoitspan pipes were discovered, followed by those at Bultfontein, De Beers and Kimberley Mine (now closed) in 1871, Koffiefontein in 1880 and Wesselton in 1891. Finsch Mine was a more recent discovery in 1961. Finsch, due to go underground in 1986, and Koffiefontein, are fully self-contained with their own treatment plants, housing, services etc. The four Kimberley mines deliver ore by conveyor to a central treatment plant and use joint ser- vices. ,+lining is presently carried out mainly by Block Caving which has replaced the earlier Chambering method. De Beers Mine, on a limited scale, and Wesselton Mine have recently started to use Sub-Level Caving, and both Bultfontein and Du- toitspan are scheduled to use the same method at a later date. The combined ore tonnage produced by the Kimberley Mines is 17 400 tons per day, Koffiefontein 12 000 tons per day and Finsch Mine 17 000 tons per day. THE EVOLUTION OF MINING METHODS Initially, the open-cast mining was carried out to as great a depth as safety would permit. In the early stages of open-cast mining, the Kimberlite was removed in horizontal slices from levels 13m to 20m apart. The rock broken in the workings was raised to the surface by means of endless rope haulage inclines. As lower levels were attained these rope haulages were replaced by aerial hoists, inclined shaft haulages or inclined skip-ways. These methods were dependent upon the strength of the unsupported pipe walls, which became increasingly unstable with depth and a system was adopted whereby the ground mined in open excavations was tipped down passes to tunnels connecting with vertical hoisting shafts sunk in the country rock. As work in the open pit became more and more impracticable due to extensive falls of ground, this system was replaced by a combination of open-cast and underground mining, and eventually by wholly underground methods. Chambering With the advent of underground mining, a system known as "Chambering" (or "Inclined Chambering") was evolved. It was essentially a combination of shrinkage stoping and caving, whereby advantage was taken of the high pressure exerted by the loose rock that had accumulated in the open excavations due to the collapse of the pipe walls. This mining method was introduced around 1890 and was successfully employed, with little change, up to the early nineteen seventies. In this system each mine was served by a main hoisting shaft situated about 300m in the country rock. Main haulage cross cuts were driven from the shaft to the pipe at 200m (600ft) vertical intervals. A sub-vertical shaft, situated close to the pipe provided access to the working levels. Chamber levels were established from the sub-vertical shaft at 13m (40ft) intervals. See figure 2. On each of these levels successive 6,8m wide “cuts" were mined out across the pipe, from one rock contact to the other, and extended vertically to the footwall of the level above. Every 6,8m cut was made up of a chamber and a "back-pillart1, each approximately 3,401 wide. The chamber was mined by overhand shrinkage stoping. The levels were carried "en echelon", with each chamber on a level approximately three cuts ahead of the chamber on the level

Jan 1, 1981

By R. F. Sheldon

The discovery of the Carlin deposit was the result of discriminating geologic research and prospecting devoted to the objective of finding a gold deposit that could be mined by open pit methods. By the late 1950s, Newmont Mining Corp. was becoming increasingly concerned about the trend of rapidly rising underground mining costs unless bulk mining methods could be applied. The desirability of finding an open-pittable gold deposit was apparent. The attention of Newmont geologists was directed to Nevada by the publication of two papers: "Paleozoic rocks in North-central Nevada" (1958) and "Alinements of Mining Districts in North-central Nevada" (1960), both authored or co-authored by Ralph J. Roberts of the United States Geological Survey. John Livermore, geologist of Newmont Exploration Ltd., attended a talk given by Roberts in Ely, Nevada in the summer of 1961. These papers and the address by Roberts gave details concerning the Roberts Mountain overthrust, a 483 km (300-mile) long, shallowdipping fault which pushed clastic and volcanic rocks of early and middle Paleozoic age eastward over younger marine formations. Subsequent uplift and doming with consequent erosion had locally removed the upper plate strata exposing windows of lower plate carbonate rocks. Roberts noted that the principal mineral deposits of the region, including gold, were associated with these windows which exhibited a preferred alinement. In the summer and fall of 1961 Newmont's exploration geologists John Livermore and J. Alan Coope began a systematic examination of gold occurrences associated with these windows, aided by further discussion with geologists of the US Geological Survey. Attention was directed to the Lynn and Carlin windows, particularly to those formations lying immediately above and below the Roberts Mountain fault. It was appreciated at the time that many of the gold occurrences in the region were unusual in that no colours were obtained even when panning high grade samples. A large number of rock samples were collected and analysed using fire assaying procedures, and background values for gold were established. This systematic rock sampling resulted in the identification of a distinct area of anomalous gold values, and a block of claims was staked in late October, 1961. These claims, plus an adjoining optioned 32.4 ha (80 acres) of ground, cover the main area of the present Carlin gold mine. Prior to snowfall that winter, one of the bulldozed assessment pits required to establish a claim's discovery exposed 24 m (80 ft) of mineralization assaying 0.007 kg/t (0.22 oz per st) gold. In the spring of 1962 a program of trenching, sampling, and geological mapping followed by rotary drilling was underway. Other properties in the area were acquired. The generally undistinguished nature of the dolomitic siltstones and silty dolomitic limestones hosting the micron-sized gold particles, coupled with the lack of visually associated guide minerals, made identification of the gold bearing areas very difficult. The entire drill column had to be assayed to ensure that values were not overlooked, as very often sections that might be assumed to be waste turned out to be high grade. On September 10, 1962 a high grade intersection of 24 m (80 ft) assaying over 0.03 kg/t (1 oz per st) gold was encountered in the third hole drilled. An expanded program of both rotary and diamond drilling led to the further delineation of the ore body. By December 1963, the exploration program had established an initial reserve of 10 Mt (11 million st) grading 0.01 kg/t (0.32 oz per st) gold. A 1.8 kt/d (2,000 stpd) processing plant was constructed and the first gold bullion was poured in May, 1965, just two years and eight months after the discovery hole was drilled. REFERENCES Roberts, R.J., Holz, P.E., Gilluly, J., and Ferguson, H.G., 1958, "Paleozoic Rocks in North-central Nevada," Bulletin American Association of Petroleum Geologists, Vol. 42, No. 12. Roberts, R.J., 1960, "Alinements of Mining Districts in North-central Nevada," Professional Paper 400-B, US Geological Survey, Article 9.

Jan 1, 1985

By T. C. Poulter, B. M. Caldwell, D. E. Brink

The flash X-ray has been used more than a decade to study the configuration of the jet from a shaped charge. The high-speed, rotating-mirror smear camera has provided time-distance graphs of detonations and shock fronts in transparent materials, and the high-speed framing camera has given pictorial representations of the progress of explosive phenomena. Adequate means for measuring the paths and velocities of all parts of the shaped-charge cavity Iiner during the collapse phase have not existed heretofore. A technique enabling the single-lens framing camera to make stereoscopic photographs of the cavity liner while it is collapsing has been developed. Analysis of this photographic record gives the directions and velocities of various parts of the Iiner surface, permitting direct quantitative measurement where this has previously been impossible. Significant improvements in shnped-clrarge design are expected to result. INTRODUCTION The familiar oilwell jet perforating charge, generally referred to as a "shaped charge", is related to the Mun-roe charge first described by C. E. Munroe in 1888 and later by Neumann in 1910. It differs in construction from the Munroe charge in that its cavity is lined with some inert material (usually metal)', and it differs in performance by projecting a fast jet of dense liner material against the target instead of a stream of expanding detonation products. Fig. 1 represents a typical shaped charge in axial cross section. The charge is circularly symmetrical about its longitudinal axis. The cavity at the right is lined with metal, usually copper; and, with the exception of air, the liner is otherwise completely empty. High-explosive material is intimately in contact with the convex surface of the liner and extends in the other direction to the booster. This is a pellet of high explosive having somewhat different characteristics and serves to couple the detonator to the main charge. When the detonator is fired, a small hemispherical reaction zone of very high temperature and pressure occurs at its output end. The booster responds, adding its own latent energy to the reaction, thereby initiating the main charge. The reaction progresses throughout the main charge at a velocity of 7 to 8 mm/microsecond until all the explosive material has reacted. With a reaction pressure of about 160,000 atm, the surrounding material is moved rapidly away from the high-pressure zone. As the detonation proceeds along the cavity liner, moving from apex to edge, the liner is collapsed toward the axis of the cavity, where liner material from all parts of the concave surface converges. A zone of extraordinarily high pressure results. with only one escape path—outward along the axis. The particles from the liner, thus, become a jet of very high velocity, with a density characteristic of the material from which the liner was made. The kinetic energy of the jet, concentrated in a small area of the target, has great penetrating power. It overcomes the strength of any target, and the target material crumbles or deforms plastically to escape from the path of the jet. An unlined cavity charge, on the other hand, projects a turbulent body of hot gaseous material against a target and erodes the exposed surface. With equal quantities of explosive, the penetration effectiveness of an unlined charge is a small fraction of that of a lined charge. Given an explosive charge with a detonator at one end and an axially symmetrical metal-lined cavity at the other, one has a perforator of considerable potency. To maximize the penetrating and other desirable effects,