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By A. S. Rhea

SALT-WATER disposal in the East Texas field by means of injection into the Woodbine sand is a relatively new procedure. The purpose of this paper is to present what has been accomplished to date in this field, with-out detailed discussion of other means of disposal, such as surface impound-ing for evaporation or release into surface streams. Subsurface disposal of salt water is still in an experimental stage in the East Texas field, and, though water can be injected into the Woodbine sand, the economic problem is not as yet solved. As past experience has been that fields producing from the Woodbine formation produce large quantities of salt water, it was assumed early in development of East Texas that it would be necessary to dispose of a tremendous quantity of salt water in the later periods of production. There has been a gradual increase in the amount of salt water pro-duced in the field, until at present it is about 200,000 bbl. per day from approximately 4400 wells. It is only natural to expect an increase in the number of wells producing water, as well as the amount of water produced in the future, and it has been estimated that if present practices con-tinue the production of salt water may ultimately be as much as 700,-000 bbl. per day. The Woodbine sand is found in the East Texas field at an average depth of 3650 ft. The oil-bearing portion of the sand lies between the depths of 3050 ft. below sea level, on the east edge of the field, and 3320 ft. below sea level on the west, which depth was the estimated original average oil-water contact. The Woodbine sand under the field has the form of a broad tilted wedge pinching out on the eastern side of the field and dipping gently into the East Texas Basin toward the west. The Woodbine sand in the East Texas field has an average porosity of approximately 25 per cent and is highly permeable, the average being about 1300 millidarcys.

Jan 1, 1940

By J. S. Levine, M. Prats

Several methods are available for calculating the performance of solution-gas-drive reservoirs from the PVT properties of the oil and from the relative permeability and other properties of the formation. These methods require a number of simplifying assumptions. The present method of computation has made use of a high-speed computer to solve simultaneously the nonlinear partial differential equations that describe two-phase flow by solution-gas drive in order to calculate the performance of a reservoir. Some of the results obtained by the nonlinear partial differential equation solution are compared with those obtained with an approximate method, which has been called the semisteady-state solution. The pressure and saturation profiles from the wellbore to outer boundary calculated by the two methods are compared for one constant-terminal-rate case and two constant-teminal-pressure cases. The agreement in these profiles, as well as in the values of average reservoir pressure and cumulative recovery, leads to the conclusion that, for most engineering calculations, the semisteady-state method will give a reasonable approximation to the numerical solution of the differential equations describing solution-gas drive. An unfavorable (as regards ultimate oil production) set of relative permeability curves was used in the calculations in the belief that the effect of the parameters which were studied would be emphasized to a greater degree. Furthermore, the reservoir was assumed to be completely homogeneous, and these results should not be considered applicable to any other type of reservoir. Gravity effects are not considered. The absolute permeability was varied from 25 to 0.5 md. At an economic limit of 2 B/D, the recovery for a 25-md reservoir is about 1.8 times as great as that for a 0.5-md reservoir. The effect of permeability on the producing gas-oil ratio is minor. Once PVT properties of the oil and the relative penneability properties of the reservoir are fixed, the pro- during gas-oil ratio is found to be a function of the fraction of oil recovered. Well spacings of 10, 40 and 80 acres were considered. For the assumed homogeneous-reservoir properties, the effect of spacing on recovery at an economic limit of 2 B/D was very slight. Certain dimensionless groups can be used to extend the results to other fields having different penneabilities, spacings, reservoir thicknesses, well radii and porosities, so long as the PVT and relative permeability properties are similar to those used in this paper. INTRODUCTION An important aspect of reservoir engineering is the prediction of the performance of the reservoir based on the limited information normally available early in the life of a field. Usually, soon after a field has been discovered, it is necessary to decide upon the spacing to be employed and the production program to be used for most efficient utilization of reservoir energy. Where these decisions have not been dictated by legal or political considerations, they have frequently been based on experience factors which indicate that two fields with similar characteristics will probably have similar performances. To the extent that the industry has been successful in economically exploiting oil reservoirs, this rule-of-thumb method has been found to have merit. A large amount of theory has been developed with which it is possible to predict the performance of a reservoir by using certain known properties of the oil and the formation. Because of the difficulties of the mathematics involved for a drainage area having square or rectangular boundaries, the problem usually has been simplified and idealized by assuming each drainage area to have radial symmetry and the field to be represented by a number of these drainage areas. Muskatl was one of the first to formulate the theory for two-phase flow and to solve the equations for a few cases by numerical integration, using the relatively slow means of computing available at that time. His equations describing

By Adrian C. Dorenfeld, John T. Oleksy

Fuel problems of two lesser developed countries (LDC 's) are presented. In particular, the paper reviews cement technology and the fuel-intensive cement industries of the Philippines and South Korea. Based on current financial and cost data, analyses are presented of the proposed substitution of high-sulfur and slow-sulfur American coal in place of fuel oil. The analyses show that economic substitution of coal for oil is not only based on relative prices for coal and oil, as expected, but also to be an inverse function of degree of intensity of energy use, and a direct function of financial incentives such as accelerated depreciation, taxes on the savings made by the use of coal, and cost of capital.

Jan 4, 1979

NORTH AMERICA Number Members Alaska 43 Canada 340 Mexico 205 Newfoundland 4 United States Alabama 39 Arizona 120 Arkansas 11 California 699 Colorado 168 Connecticut 89 Delaware 18 District of Columbia 140 Florida 20 Georgia 6 Idaho 78 Illinois 214 Indiana 77 Iowa 12 Kansas 24 Kentucky 33

Jan 1, 1936

By C. E. Birchenall, J. R. Stubbles

ThE equilibrium between lead, lead sulfide (PbS), and atmospheres of H2 and H2S has been determined previously by Jellinek and Deubel,' and SudO.' Blbem and Kroger3,4 used Jellinek's data to calculate the sulfur dissociation pressures of PbS crystals (of

Jan 1, 1960

By Isaac H. Odell, Arthur W. Hixson, Lincoln T. Work

Although a few engineers have recognized the problem of the mechanism of filtration it has never been studied in a quantitative way. A background for a better understanding will be afforded by a survey of certain fundamental principles of filtration. In recent years, authorities have become practically unanimous in saying that the actual filtering should be done, not by the cloth itself, but by a layer of solid particles deposited upon it. Thus, a cloth may be used in which the openings are considerably larger than the particles to be filtered, provided that groups of particles form an arch or bridge over each opening. This can be accomplished if the initial pressure is so low that the arch will not be broken while it is still thin. According to Wright,' the chief advantage of this method of filtration is that the open-weave cloth does not become clogged as easily as does one of closer weave, where the fine particles often work their way into the interior of the cloth and become caught there. If the character of the material to be filtered is such that it will not form a good film, then it becomes advantageous to use "filter aids" for pre-coating the cloth with a desirable layer of solids, and the coating need be only of a very slight thickness for good results. Until the film is formed, the first runnings of the filtrate may be cloudy, often necessitating refiltering this portion. A factor which has considerable effect on the performance of filters is the concentration of solids in the pulp. According to Wright, a high concentration of solids gives a porous cake, since the flow is then slower than with a thin pulp. The porosity of the cake is also affected by the pressure: a high initial pressure tends to compact the cake, and the same effect results from fluctuations of pressure. In his studies of the relation between pore size and particle size, Hatschek2 observed that the particles in a pulp tend to arrange themselves with maximum cross-section parallel to the direction of flow and that this is more marked at the higher vclocfiies. Thus, at high pressure and high velocity, a certain pulp

Jan 1, 1926

By D. G. Westlake

THE structure observed in thin foils by transmission electron microscopy is not necessarily representative of the initial bulk material. We have shown that hydrogen, originally present in a specimen of vanadium, niobium, or tantalum, can be concentrated by the thinning operation used to prepare foils from the bulk materials.' These foils exhibited domain walls indicative of a tetragonal distortion of the bcc unit cells, probably caused by ordering of the hydrogen in preferred interstitial sites. We believe this ordering does not occur at room temperature in the bulk material which contains less than 1 at. pct H. In a continued effort to prepare metal foils whose hydrogen concentrations remain constant during preparation, we have developed some new chemical and electrothinning techniques for vanadium. Each of the procedures presented in Table I produced foils that were thin enough for transmission electron microscopy. The experiments described below were performed to test the effect of each thinning procedure on hydro-

Jan 1, 1968

By W. B. Mather

A SURVEY of the technical literature concerned with oil well drilling methods and particularly with rate of penetration by various cutting media on different types of rock provides a mass of conflicting data. This is especially true with respect to, (1) the relative efficiency of diamond, rotary and percussion drilling operations; (2) the costs of the various types of drilling; and (3) the hardness factor of various types of rock and cutting media. The latter is the particular phase of the subject discussed here. The cause of the confusion may be attributed partially to a lack of 'basic knowledge concerning the physical properties of the cutting media and of individual types of rock. It is also due in part to a general lack of uniform standards in the industry for evaluating the hardness versus drillability of rocks while possible errors in interpretation by the authors of the various papers pertaining to the subject may be an additional cause of misunderstanding. The hardness of a mineral is defined by Dana' as "the resistance which a smooth surface offers to abrasion." Webster' states that hardness is "the cohesion of the particles on the surface of a body (as a mineral) as determined by its capacity to scratch another or be itself scratched." Rock hardness, although not usually defined in publications describing rocks, is assumed to be the same as the minerals which compose the rock or the resistance of the rock to abrasion. Drillability of a rock may be defined arbitrarily as the rate at which rock is penetrated by a definite type of drilling tool with the assistance of a specific type of cutting surface or medium and under speci- fied conditions as to speed of rotation and applications of pressure. Eventually the term penetrability may have to be employed to include any method of making boreholes in rocks, since methods other than those using conventional cutting surfaces are already in operation. A case in point is the jet piercing methoda which employs an oxy-acetylene flame as a means of penetration. The word hardness as used by the drilling fraternity has a variety of meanings which depend upon the type of drilling method employed. For example, in diamond drilling, hardness is interpretated as resistance to abrasion. In percussion drilling the term implies resistance to impact or indentation, while in rotary drilling, hardness is considered anal-agous to compressive strength. The three types of hardness—abrasion, impact and compression—are conflicting variables for different types of rock. Nevertheless, the term hardness is applied indiscriminately by the industry for the resistance of rock to penetration by any type of drilling technique. Abrasion Hardness Since 1818, Moh's scale of hardness, although only relative in nature, has been accepted universally as the method of measuring mineral and therefore rock hardness. In 1933, Ridgway, Ballard and

Jan 1, 1952

By A. S. Richardson

The design of mine airways receives, in general, very little engineering treatment. To a large extent this is, of course, due to the fact that information upon which to base calculations is seldom available in definite form. In many rnines the extent of the workable deposits may not be known at the time an air course is planned, and the length of service is indeterminable. Volume of air required for ventilation rnay also increase with extension of the life of a property, and natural obstacles, such as heavy ground, often place limitations upon the size of opening that can be maintained. Perhaps most of the shafts used in the ventilation of metal mines were originally intended as hoisting shafts, and many of the tunnels and crosscuts similarly used, were also designed as haulage ways or for other operating purposes. Even when a shaft is sunk solely for ventilation, custom usually dictates a size and shape consistent with the type commonly used for ordinary operations, because experience has demonstrated it can be excavated and maintained. In fact, most ventilation openings were never made with any consideration as to economic design for such use. When an air course is required to pass air between different parts of mine workings under a fixed pressure, the size of opening required for a given volume of flow, is, of course, determined by the resistance factor of the air course. It is then impossible to proportion the opening with reference to economics of power consumption, and this paper is limited to the design of openings where ventilation pressure or power consumption may he varied for economic advantage. As the size of an air course is increased, the expense involved in excavation and support is, of course, also increased, but the charge for power is decreased. In the case of a mine that has only a short assured operating life, it would not be ecouomical to provide large openings in order to reduce power bills; but when operations are likely to continue for a long time, the power charge becomes a very important economic factor. Although lack of definite information on such factors as service life and costs of excavation and support affects the design of an air course, it appears, from a mathematical analysis of the problem, that considerable variations may be made in estimating these factors, without greatly

Jan 1, 1927

By Michel Daghert, Michel David

In the exploitation of large porphyry type deposits where a cut-off has to be applied and where "internal waste" exists, planning people ask the geologist for the estimated grade of small blocks. Comparison with actual production figures and blast hole assays shows that any kind of weighted average of neighboring DDH (diamond drill hole) samples provides an excessingly smooth picture of reality, resulting most of the time in overevaluation of tons of ore. Solutions based on examination of Past production data exist14 but it would be necessary to have a solution which can be applied before production starts as the preceeding problem also results in an overestimation of the total reserves, often by as much as 20%, thus changing the mine life accordingly and destroying any "optimum planning" which could have been done. The proposed alternative is to forecast the grade distribution of small blocks in a larger block, sufficiently well documented. Several Possible ways to estimate these distributions are reviewed and emphasis is placed on recently developed methods (source G. Matheron) where the coefficients of the orthogonal polynomial series expansion of the distribution of any size of block are directly estimated by kriging. The efficiency of the technique is tested on a level of the Cyprus Pima open pit where the predicted recovery is compared to actual production figures.

Jan 1, 1980

By C. L. Prokop

A laboratory investigation has been made of the effects of mud hydraulics upon the formation and erosion of mud filter cakes. The tests were conducted to simulate drilling conditions as nearly as possible. The formation of mud filter cake in a drilling well does not proceed at a uniform and unbroken rate. Instead, the rate of cake accumulation depends upon whether or not the mud is being circulated. If the mud column is quiescent, filter cake formation is a smooth function of the filtration characteristics of the system. If the mud is being circulated filter cake formation depends not only upon the filtration characteristics of the mud but also upon the erosive action of the flowing mud column Filter cakes formed during continuous mud circulation were observed to reach an equilibrium thickness after several hours' circulation. Mud circulation was maintained at a constant volumetric rate throughout each experiment. The fluid velocity at equilibrium cake thickness was dependent upon the thickness of the filter cake. Muds having exceptionally high water loss deposited thick filter cakes in spite of very high eroding velocities. The muds having good filtration characteristics deposited thin filter cakes at equilibrium circulating velocities well within tile range of those in a drilling well. It was observed that filter cakes deposited during stagnant filtration were quite difficult to erode by mud circulation. The - rate of crosion computed from the rate of filtrate accumulation after equilibrium cake thickness had been reached was in reasonable agreement with the rate of erosion obtained by direct observation. Continuous mud circulation usually caused the permeability of the filter cake to decrease with time. INTRODUCTION Many of the difficulties encountered during tile drilling of a well have been attributed to the loss of water from the mud and the attendant deposition of solids upon the walls of the hole. Past experience has shown that a reduction of the filtration rate of the drilling fluid eliminates or greatly reduces these difficulties. Definite filtration requirements, however, are hard to establish for a given set of conditions. This is due. in part, to the fact that the usual filtration test performed upon mud doe? not simulate well conditions as closely as desirable. The filtration characteristics of a mud are customarily determined by means of the standard low-pressure API wall-building tester.' In this instrument a filter cake is deposited upon a horizontal bed under a pressure differential of 100 psi. The rnud is quiescent during the filtration period. In actual practice. mud filtration occurs within a well under quite different conditions. One of the major differences is that mud flows upward across the filter bed as the filter cake forms. This undoubtedly produces a change in the filter cake which cannot be reflected in the results of the API test. The laboratory work described in this paper had as its primary objective a better understanding of the influence of mud circulation upon the thickness and ,characteristics of the filter cakes deposited under conditions similar to those existing in a drilling well. ANALYSIS OF PROBLEM Once a permeable formation is penetrated by the bit, filtrate from the mud flows into the formation. 'he mud solids plaster against the walls of the hole, forming a filter cake. If the mud column is stagnant, that is, if it is not being circulated. the filter cake will increase in thickness until the hole is filled. Prior to the time that the hole is filled, the thickness of filter cake existing at any given time will be a function of the filtration characteristics of the mud, the temperature, and the pressure differential. The effects of these variables have been investigated in the past for both flat bed filtration2'3 and for radial filtration.' When the mud is circulated in a hole in which a filter cake i. being deposited. some of the solids that would ordinarily deposit in the filter cake will be carried away by the eroding action of the mud. This will limit. filter cake thickness. Some work has been done to determine the effect of flow upon the filtration rate in a circulating mud system' but little work has been done upon the factors which determine the filter cake thickness existing in a circulating system. On first sight it would appear that the major factors controlling filter cake formation in a circulating system should be: 1. The rate of deposition of solids from the mud. 2. The erosive force that the flowing mud exerts upon the filter cake. 'A. The erodabilitv of the filter cake. 4. Any change in filter cake characteristics attributable to the scouring action of the mud. The rate at which solids are deposited from the mud will be controlled to a large degree by the filtration characteristics of the mud, the pressure differential. the temperature under

Jan 1, 1952

By George R. Gardner, Kenneth B. Ray

THE application of wet cleaning processes for the beneficiation of bituminous coal has created in some localities a problem in the recovery and disposal of fine solids in the washery water. The maximum allowable concentration of these solids in the washery circuit that is consistent with satisfactory operation of the cleaning units remains a controversial point. It has, however, been generally recognized for a number of years that some degree of control should be exerted over the concentration of such material. Early methods of control involved intermittent or continuous discharge of a portion of the plant slurry into near-by streams or settling ponds, and this persisted even after the adoption of settling tanks or cones. The practice is not always economical, for two reasons: first, because of the waste of coal, and, second, because it is sometimes difficult and expensive to obtain an adequate supply of water. The latter condition often is aggravated by chemical treatment of the make-up water that is necessary to prevent excessive corrosion of plant equipment. The modern approach to this problem has been the adoption of continuous thickening devices. However, because such equipment occupies considerable plant space and is somewhat expensive to construct, the practice of loading these thickeners beyond their rated capacity is widespread, therefore often the performance of such units is not altogether satisfactory. In recent years, European washeries have increased the rate of settling of the fine coal from the slurry by the addition of specially prepared starch solutions. These reagents cause flocculation of the solid particles and thereby increase the rate of sedimentation. The first important work on the flocculation and clarification of coal slurry was done by Henry, who developed the Henry process and used it successfully in Belgian preparation plants. Soon after the initiation in Belgium, the process was introduced into England, where subsequent investigation developed similar processes, which were successfully used in

Jan 1, 1939

By I. R. Kramer

The apparent activation energy, U, JOY plastic deformation of aluminium, copper, and gold single crystals and polycrystalline aluminum was found to he dependent upon surface condition. At low strains, the value for U for aluminum single crystals was 4200 cal per mole ; however, this value decreased to 920 cal per mole when the specimens were polished during the test at a rate of 50 X 10-5 win-1 The activated volume also was found to increase with the rate of metal removal. That the surface plays an important role in plastic-deformation processes has been adequately demonstrated.1-4 Of particular interest at this time are the observations4 that the creep rate of single crystals of fee metals is increased markedly when specimens are polished during the deformation process. These observations were interpreted in terms of a dislocation debris layer at the surface regions of the crystal. This layer, which appears to be a region containing a high concentration of dislocations or segments of dislocations, not only will impede the egress of dislocations on the primary slip system, but will also provide a barrier against which dislocations may pile up. By this means, a stress which opposes the motion of dislocations will be generated. It is considered that this stress must be taken into account in any theory which describes the plastic-deformation process. Although a detailed dislocation model cannot be given at this time, it may be stated that the stresses opposing the motion of dislocations arise from the dislocations in the debris layer, Ts, and ti due to the resistance to motion imposed by internal obstacles such as Lomer-Cottrell barriers, subgrain and grain boundaries, forest dislocations, and so forth, as well as the resistance due to the presence of jogs on dislocations and the interaction of the stress fields of dislocations, and so forth. In essence, we have added an additional term, ts, to the usual form of the equation T = Ta- Tb, where Tb is the back stress due to the various types of obstacles which impede the motion of dislocation. Accordingly, the net stress, t, acting on a dislocation may be written as where Ta is the applied shear stress. For deformation which occurs by a thermally activated process5,8 ?=Ae-u/kT = A e(u,-VT)/kT [2] and U = Uo- Vt PI where U = activation energy for plastic flow, y = shear strain rate, Uo = activated energy for plastic deformation in the absence of stress and depends upon the particular deformation process, V = activated volume, k = Boltzmann constant, T = temperature, OK, and A = frequency factor. From Eqs. [I] and [3], U = Uo- V(Ta - Ti - TS) [4] and it is seen that if Ts decreases, the activation energy, U, must decrease if V remains constant. However, as will be shown, V is a function of the surface conditions. In their work on aluminum, Lytton and coworkers7 reported that at low strains the activation energy was 3400 cal per mole and at high strains it was 28,000 cal per mole. The lower value was interpreted in terms of the energy necessary to overcome the Peierls force, while the higher value was associated with the energy necessary for dislocations to move by a cross-slip process. It will be shown that the lower activation energy is associated with the surface debris layer rather than with the Peierls force. EXPERIMENTAL PROCEDURE Single crystals having a nominal cross section of 1/8 in. and a 3-in. gage length were prepared by a modified Bridgman technique using a multiple-cavity graphite mold. The initial purity of the metals was 99.997, 99.999, and 99.999 pet for aluminum, copper, and gold, respectively. The orientation of the specimens is given in Fig. 1. The measurement of the activation energy as a

Jan 1, 1964

By Wolfgang O. Gluschke

INTRODUCTION Virtually all countries have general legislation covering most aspects of mining and mineral processing, including investment and tax laws, safety and health regulation, and specific rules applying to exploration reporting procedures. Yet most developing countries negotiate individual agreements with foreign investors for major mineral projects, in particular when all or part of the expected output is exported. Such agreements are known under several different names, as for example concession agreement; technical assistance agreement or service contract; management agreement; production sharing contract; joint venture; or contract of work. The objectives and conditions governing such contracts are as diverse as these titles are. While a mining project is only one among a large mumber of similar undertakings in industrial countries, in most developing countries such projects usually have a dominating position within the country's economy. The general legislation is often inadequate and insufficiently detailed to cover all aspects of large projects with extensive foreign inputs and international markets. Therefore, the need for specific arrangements. Specific arrangements may also become necessary due to the complexity of markets for mineral products, the large number of participants - equity partners, suppliers, financial institutions, buyers, etc. - or economic and political conditions in the host country and fiscal regulations in the home country of the investors. Most agreements are concluded between governments, or government agencies, in developing countries and private industry headquartered in developed market economies - Canada and the United States of America in North America, Western European countries, Japan, South Africa and Australia - or, in some instances, with state-owned firms in these countries. There are fewer agreements between developing countries and the centrally planned economy countries - Eastern Europe and the USSR - which cannot be considered here because of absence of adequate information. Agreements can be concluded at any of the different stages from exploration through closing of a mine. While most contracts are negotiated at the time when a potentially viable ore body has been delineated and investigated sufficiently to anticipate a realistic time for commencement of production, other agreements start with "grass-roots" exploration, during operation of a mine, smelter or refinery or, exceptionally, after a temporary shutdown. While several major provisions are included in all types of agreements, others are covered in much more detail (or not at all) in some contracts reflecting the specific work programme of each project. While it is realized that subdivision and classification of complex contractual relations is necessarily arbitrary to a certain extent, such grouping is useful in practical terms, especially for the non-specialist. The following discussion, therefore, will start with a description of the major forms of contractual relation - concession, investment contract, joint venture, management agreement, service and production sharing contracts. Major issues covered in all, or almost all, agreements will then be dealt with. These include stages of development, corporate structure and control, fiscal arrangements, infrastructure development and environmental law, and dispute settlement and renegotiation. FORMS OF CONTRACT RELATIONS There has been marked change in the type of agreement entered into during the 1960s (and before) and the early 1970s and the period thereafter. While straight "concessions" - granting of exclusive rights to foreign investors over a certain deposit or areas with little restriction or well-defined obligations - were the rule during the first half of this century, more complex agreements have evolved since. Some of these cover every conceivable aspect, foreseen or only possible, and define rights and obligations of all partners to the last detail. Over the years, agreements have without doubt become more beneficial for the developing host countries, in part as a result of well-publicized nationalizations, as for example in Chile, and in part due to changed attitudes in all countries

Jan 1, 1985

By S. FROES ABRUE

No new oil fields were discovered in Brazil during 1945. Production for the year reached a total of 79.329 bbl., all coming from the four fields in the Baia basin; the Lobato-Joanes field produced 6728 bbl.; Aratu, 6717 bbl.; Candeias, 36,663 bbl. and Itaparica, 29,211 bbl. The total cumulative production from the Baia basin from 1939 to 1945 inclusive amounts to 210,000 bbl., all of which is utilized locally.

Jan 1, 1946

By M. J. Richards, A. Szirmae, G. R. Speich

The formation of austenite from ferrite, ferrite plus retastable carbide, spheroidite, and pearlite has been studied in a series of irons, Fe-C alloys, and plain-carbon steels using fast heating techniques. In the absence of carbide, austenite nucleates at ferrite/ferrite grain boundaries; nucleation is followed by the rapid growth characteristic of a massive transfornation. The trarnsformation occurs at 950°C at heating rates of 106º C per sec and cannot be suppressed. Metastable carbide dissolves before austenite forms and does not influence the transformation kinetics. For spheroidite structures, austenite nucleates preferentially at the jinction between carbides and ferrite grain boundaries. Growth from these centers proceeds until the carbide is completely enveloped; subsequent growth occurs by carbon diffusion through the austenite envelope. For pearlite structures, austenite nucleates preferentially at pearlite colony intersections. Carbide la)?zellae dissolve at the advancing austenite interface but complete solution of carbide does not occur; the residtial carbide is eventually dissolvled or spheroid-ized depending on the carbon cuntent. The magnitude and temperature dependence of the austenite growth rate into Fe-C pearlite when incomplete carbide dissolution is assumed are satisfactorily explained by an approximate colume diffusion model. The impurities present in plain-carbon steel reduce the growth rate of austenite in comparison to that jound in an Fe-C alloy. The formation of austenite has been studied in much less detail than the decomposition of austenite. This is primarily a result of the importance of harden-ability in determining the mechanical properties of steel. Recently, more interest in the kinetics of austenite formation has resulted from the discovery by Grange1 that rapid heating techniques strengthen steel by refining the austenite grain size. Although the strengthening effect is not large, it is accompanied by no loss in ductility. In addition, interest continues in rapid heat treatment of low-carbon steel sheet for tin plate applications.2,3 Among the few systematic studies of austenite formation are the early work of Roberts and Mehl4 on formation of austenite from pearlite and recent work of Molinder5 and of Judd and paxton6 on formation of austenite from spheroidite. Also, Boedtker and Duwez7 and Haworth and paar8 have recently studied the formation of austenite from ferrite in relatively pure iron, Kidin et al.9,10 have studied the formation of austenite in 8 pct Cr steels, and Paxton has recently discussed various aspects of austenite formation in steels." The present work was undertaken to determine the kinetics of austenite formation for a variety of starting structures including ferrite, ferrite plus metastable carbide, ferrite plus spheroidal cementite, and ferrite plus pearlitic cementite. Emphasis was placed on determining the active sites for austenite nucleation, determining the temperature and time range of austenite formation, and in the case of pearlite a careful study of the growth rate of austenite was made in the absence and presence of impurities. By using a variety of heating techniques including laser-pulse heating, it has been possible to study austenite formation in an isothermal fashion over a wide range of temperatures. EXPERIMENTAL PROCEDURE The alloys studied in the present work are a zone-refined iron with 4 pprn C, an Fe-C alloy with 130 pprn C, 2 Fe-C alloys with 0.77 and 0.96 wt pct C, and a plain carbon steel with 0.96 wt pct C. The zone-refined iron and Fe-C alloys contained 60 pprn and 200 pprn total substitutional impurities, respectively. The plain carbon steel contained 2400 pprn Si, 2000 pprn Mn, and 900 pprn Cr. Various heat treatments were given to these alloys to produce different starting structures of equiaxed ferrite, ferrite plus metastable carbide, fine pearlite, and spheroidite. These heat treatments are given in Table I. A wide range of heating rates were employed in this work because many of the reactions occur so quickly at temperatures in the austenite range that they are completed during the initial heating cycle unless very fast heating rates are used. Essentially the same heating techniques employed by Speich et a1.12 and Speich and Fisher13 were used in this work. For time intervals of 2 sec to 20 hr, simple hand immersion of 0.010-in. thick specimens in a Pb-Bi bath was employed. These specimens were quenched in a 10 pct NaC1, 2 pct NaOH aqueous bath. For time intervals of 100 m-sec to 2 sec, an automatic dunking and quenching device was employed with 0.002-in. thick specimens. Again, liquid Pb-Bi baths were used for a heating medium but now helium gas quenching was employed. For time intervals of 2 to 100 m-sec a laser heating device was employed with 0.002-in. thick specimens; a helium plus fine water-droplet spray was now used for quenching. Additional information on heating times shorter than 2 m-sec was obtained by study of the zones around the centrally heated laser spot. Here diffusion of heat from the centrally heated zone raises the temperature of the specimen locally to all temperatures between ambient and the peak temperature, but for times of the order of microseconds. All the heat-treated specimens were examined by

Jan 1, 1970

By K. R. Carson, J. Weertman

An attempt is made to ascertain the mechanism of tangle and cell formation and its dependence upon dislocation-interstitial carbon interactions. The strain-hardening behavior of single crystals of 3.25 pct Si-Fe was determined at 300° and 425°K and under conditions of both continuous and interrupted tensile strain. Significantly enhanced hardening was observed in crystals deformed at the elevated temperature, and it was further accentuated by interrupted straining. Transmission electron microscopy was used to study the resultant dislocation structures. Strain aging was found to aid tangle and cell formation at 425°K, but at both temperatures embryo tangles formed solely from primary glide dislocations, presumably by a process involving cross slip and "mushrooming". IN the course of plastic deformation all bcc metals and alloys develop a dislocation structure characterized by loose-knit groups of tangled dislocations. With increasing strain the tangles become more tightly knit and grow larger; finally a three-dimensional cellular substructure is formed:1 This process has been observed with the transmission electron microscope.'-l7 However, most investigations were confined to the study of nearly pure polycrystalline metals at relatively low temperatures. At intermediate temperatures, 0.17 to 0.14 Tm where T, is the melting temperature in degrees absolute, the mobility of interstitial impurities such as carbon is high enough to permit migration to nearby glide dislocations but is still low enough so that a significant drag force is exerted.18,19 it is also in this temperature range that a hump occurs in the curve of work-hardening rate vs temperature for iron. Analogous plots for tantalum" and columbiumzo show a definite upward trend in the work-hardening rate. Keh and Weissman1 have pointed out that this behavior may be explained solely on the basis of changes in the dislocation configuration: at low temperatures the dislocations tend to be relatively straight and uniformly distributed, but at intermediate temperatures tightly knit tangles and cellular substructure appear. The interference of these tangles with glide dislocations causes the observed increase in the work-hardening rate. This explanation appears reasonable, yet one might ask what factors cause tangle formation to be so favorable at intermediate temperatures. It seens likely that the strong dislocation-interstitial interactions which are known to occur in this temperature range are at least partly responsible," with the magnitude of the effect being proportional to the interstitial concentration. The purpose of the present work is to study the relationship between tangle formation and strain hardening in a bcc metal in the temperature range 0.17 to 0.4 Tm. Particular emphasis was placed upon a study of the effects of interstitial-dislocation interactions. Single crystals of 3.25 pct Si-Fe containing about 200 ppm of C in solid solution were used in the investigation for the following reasons: 1) The mobility of interstitial carbon in 3.25 pct Si-Fe is negligible at 300°K but increases rapidly at slightly elevated temperature22. Hence, differences between the flow curves and dislocation structures of crystals deformed at 300°K, 0.17 T,, and crystals deformed, say, at 425°K, 0.24 Tm, should be appreciable because of the enhanced dislocation-carbon interactions at the elevated temperature. This effect was accentuated in some samples by interrupted straining, thereby introducing a certain amount of aging. 2) Near room temperature, slip in suitably oriented 3.25 pct Si-Fe single crystals is largely confined to the (110) planes.23'24 Dislocation structures formed under conditions of single glide are the least complicated and their method of formation is the most easily discernable. 3) Dislocations in Si-Fe can be tightly locked with carbon atmospheres by a low-temperature aging treatment. The subsequent thinning of samples to foil thicbess causes little or no rearrangement in the dislocation structure.25 EXPERIMENTAL PROCEDURE Large single-crystal sheets of 3.25 pct Si-Fe were donated by Dr. C. G. Dunn of the General Electric Research Laboratory, Schenectady, N. Y. The orientations of the sheets were determined and slabs 1.0 by 0.25 by 0.05 in. were cut such that the desired tensile axis corresponded to the long dimension. The slabs were mechanically polished and subsequently decar-burized by heating at 1000°C for 3 days in a flowing wet-hydrogen atmosphere. A carbon content of about 200 ppm was introduced by heating at 805°C for 25 min in a flowing atmosphere of dry hydrogen containing heptane vapor. Shaped copper tools were then used to spark-machine at 0.125 by 0.50 in. gage length onto each slab. Vacuum annealing at 1225°C for 2 days followed by a quench into the cold end of the furnace to retain carbon in solid solution concluded the soecimen preparation. Continuous tensile flow curves for crystals of severa1 orientations Were obtained both at 300' and 425°K. A strain rate of 6.67 x 10-4 Per set was used in these and all other tests. Crystals oriented for single glide, B and D in Fig. 1, were subjected to a 3.5 pct plastic elongation to insure uniform slip along the gage length; they were then immediately subjected to interrupted strain cycling as indicated in Fig. 2(a). Each cycle consisted of unloading to 1.5 kg per sq mm, holding

Jan 1, 1969

The Ortiz Mountain property is located approximately 40 km (25 miles) southeast of Santa Fe and 13 km (8 miles) from the village of Cerrillos. The property embodies a Spanish mineral grant from which Gold Fields acquired about 2, 200 hectares (900 acres of surface and mineral rights. Following a period of exploration, metallurgical testing and development of a feasibility study in 1974, the decision was made, because of the higher gold price in the fall of 1978, to place the property into production. the gold is confined in a quartzite breccia; it is an unusual occurrence because the gold in this host rock is leachable at minus 19 mm (3/4 in. ). Total ore reserves are estimated at approximately 3.6 million mt (4 million st) averaging 1.89 g per mt (0.055 oz per st) based on a cut-off of 0.7 g (0.02 oz). The ore will be recovered by open pit mining at a stripping ratio of 1. 2:l. Production is scheduled for 13,600 mt (15,000 st) of ore per week or approximately 680,000 mt (750,000 st) per year. Primary crushing of pit-run material is by a 1,070 by 1,015 mm (42 in. by 40 in.) jaw crusher. Crushed ore is conveyed to stockpile and then to a 2.13 m (7 ft) standard crusher with the screened over- size crushed in a 1.5 m (5 ft) tertiary cone crusher. The screen undersize, the final product for leaching, is minus 19 mm (3/4 in. ). A 1, 360 kg (3,000 lb) sample is cut from the crushed ore stream, further crushed in a small jaw crusher and then sampled in two more stages by Vezin samplers to produce a final 13.6 kg (30 lb) sample for assaying. Pebble lime is then added to the ore on the belt as it is sent to the conveyor gantry spanning the leach pads. A tripper on this moving gantry stacks the leach ore without compaction. The entire crushing plant and conveying system is of open-air construction. Leach pads have a total length of 515 m by 60 m (1,700 ft by 200 ft) wide with usable dimensions of 490 m by 58 m (1,600 ft by 190 ft). The directional length slope is 3. 870 with 370 on the width. The pad is constructed of 175 mm (7 in. ) thick asphalt containing a rubberized asphalt membrane. It is divided by a 300 mm (1 ft) high curb into eight segments. Each segment is loaded with 13,600 mt (15,000 st) of ore, which is one week's production. Optimum working height of a pad is considered to be 3. 0 m (10 ft) with a maximum height of 4. 0 m (14 ft).

Jan 1, 1981

By John Jermain Porter

The chief purpose of this paper is to indicate the laws governing the fuel-economy of the cupola, to examine the feasibility of some of the proposals for increasing its fuel-economy, and to show that the conditions limiting fuel-economy here are altogether different from those prevailing in the blast-furnace. The two methods available for studying the thermal efficiency of the cupola and determining the factors limiting fuel-economy are: 1, by the determination of the total heat-balance; and 2, by the application of the theory of critical temperatures. The case selected as an example represents a cupola operating under fairly efficient conditions. The data are as follows : Cupola, GO in. in diameter, 15 ft. high to the charging-door, with a 9-in. lining. Bed-charge, 2,000 lb. of coke and 4,000 lb. of iron. Subsequent charges, 400 lb. of coke and 4,000 Ib. of iron. Total number of charges, 20. There was 800 lb. of coke recovered from the drop, hence the total coke burned is 8,800 lb., or 0.11 lb. of coke per Ib. of iron. Coke contains 90 per cent. of fixed carbon and 2 per cent. of moisture. 300 1b. of kindling-wood is used in lighting. 80 lb. of limestone (95 per cent. of CaCO,) is used per charge, or 0.02 lb. per Ib. of iron. Melting-loss, 4 per cent.; distributed thus : Fe, 3.5; Si, 0.25; Mn, 0.25 per cent. Average analysis of top-gases : CO2, 15.1; CO, 10.0 per cent. Average temperature of top-gases, 1,600" F. Temperature of air and stock charged, 60" F. Dew-point of air, 50 F. The items entering into the total heat-balance and their calculation are as follows: 1. Heat of combustion of fuel.—Total heat evolved = 14,580 x lb. of carbon burned + 7,200 X Ib. of wood burned. Hence B.t.u. per lb. of iron charged =

Jan 1, 1913

Clarence M. Haight, Franklin Furnace, N. J. (communication to the Secretary*).—In that part of Mr. Wilson's paper describing the Gilman cut-and-fill system, a few features do not appear to be fully explained. In what way is the waste used for filling obtained? The text mentions that if waste is encountered when driving the filling raises, these raises are stopped until the fill is needed; that seems to suggest mining the capping for fill, which might make the backs of the stopes bad as the stopes approach the top of the ore. If, on the other hand, the fill is brought to the raises in cars, are the crosscuts that are necessary in ore or waste? Is the cost of this fill included in the stoping costs given? The second, or lower, series of stopes will present a different problem to that which the first series (under the capping) gave. Although no mention is made in the text, it would seem that the second series of stopes will have to be worked on center lines which are offset half the stope width, 20 ft., from the centers of the stopes above, in order to take advantage of the "rafter" effect of the stringers so named. How is the fill to be introduced into the lower stopes? There are no crosscuts mentioned to provide for this; or is it intended to tap the filled stopes above for the fill? In the latter case the method of control of this fill will be of interest. Ore that will admit of a stoping width of 40 ft. with the lengths mentioned in the description is stronger than any with which many mining men are familiar. Has the practice of starting to mine a second stope before the preceding one is filly completed been tried to any extent? If

Jan 1, 1917