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By Giovanni Barla, Stefan H. Boshkov

This paper examines the application of gelatin mixture models to the study of factors such as gravity effects, tectonic and residual stresses, anisotropy, physical nonlinearity, time-dependent behavior of the rock mass and connectedness of the geologic space. The technique of calibrating gelatin mixtures proposed by Richards and Mark in 1966 is applied to the particular problem considered. Similitude relationships between model and prototype are discussed with reference to the constitutive equations describing the physical behavior of the rock mass. The interaction between two openings in a linearly elastic medium, under conditions of plane strain, is used as an illustrative example with the view to describing the history of stress and deformation changes on the rock medium and lining around a slusher drift while undercuts are made in block caving. The prediction of the displacement, stress field and mode of fracture associated with a given rock structure, its initial stress field, and its physical properties is essential to rock mechanics and ground control. The basic approach is to assume the rock mass to behave as a continuous, homogeneous, and isotropic medium. The classical theory of elasticity is the simplest theory based on such a concept. Specification of the elastic constants, the initial state of stress in the ground, and the boundary conditions on stress and displacement allows one to predict stress and displacement within the body. Subsequently, provided that a criterion which governs rock fracture under various stress conditions has been established, a stability analysis of the rock structure can be performed. A natural extension of the present knowledge consists of removing some of the simplifying assumptions which characterize the analytical solutions based on the classical theory of elasticity. Complex geometry, the significance of tectonic and residual stresses, anisotropy, physical nonlinearity, time-dependent behavior of the rock mass and connectedness of the geologic space are some of the important factors which are neglected in the classical approach. A better insight into the real problem can be attained with either the finite element method of stress analysis or an experimental technique. The former has already been applied with success on several occasion 1-4 and is destined to render great service in the field of rock mechanics and ground control. The latter, which will be considered in this paper, has been used extensively in applications too numerous to be enumerated. Experimental technique involves either prototype testing, if the area to be studied is accessible, or model testing. Two- and three-dimensional photo-elastic analyses are very convenient means of obtaining the stress distribution in models of complex geometry. In general, such analyses are confined to studies of stress fields around openings which are sufficiently far removed from the surface boundary so that the stresses in their neighborhood are practically equal to those produced by an initial uniform stress field; i.e., so that gravitational effects are negligible. There are, however, many rock structures where these effects must be taken into account to attain a more realistic representation of the problem. Surface and near surface rock structures, systems of openings interacting with one another, and structural components loaded only by their own weight are some of the obvious examples. This paper was written to show how factors neglected in the usual analysis based on the classical theory of elasticity, can be taken into account when gelatin mixture models are used inconjunction with photoelastic methods. The interaction between two openings in a linearly elastic medium, under conditions of plane strain, is used as an illustrative example, with a view to describing the history of stress and deformation changes on the rock medium and lining around a slusher drift while undercuts are made in block caving. The birefringence of gelatin mixtures has been known for some time. A series of experiments were carried out by Rossi in 1910,5 but after that very

Jan 1, 1970

By O. D. Gonzalez, R. A. Oriani

A thermo-osmosis technique has been used to measure the heat of transport, Q* , of hydrogen and of deuterium dissolved in a iron and in nickel, and of hydrogen in Feo.6Nio.4 in the tempevature range 400° to 600°C. For all these systems, Q* is negative and has a large temperature coefficient; an isotope effect can be established for the solutes only in nickel. The magnitude of Q* is considerably larger than the activation energy, E , for migration in the case of these isotopes in a iron, so that all variants of the Wirtz model must be rejected. The phenomenological definition of Q* is developed to show that Q* is related to the mechanisms by which the activation energy is dissipated back into the lattice, and that Q*/E is a function of the ratios of the mean free paths of electrons and of phonons to the distance of jump of the diffusing atom. A correlation is shown to exist between the sign of Q* in thermal diffusion and that of the effective charge in electro migration, and may be understood as due to the large ratio of the mean free path of electrons to that of phonons. THE study of nonisothermal diffusion in the solid state is of interest because certain aspects of the mechanism of diffusion are made manifest which remain concealed though still present in the isothermal case. The thermodynamics of irreversible processes characterizes nonisothermal diffusion, known as thermal diffusion or as the Ludwig-Soret effect, by a quantity, Q*, called the heat of transport, which essentially measures the sign and the magnitude of the steady-state concentration gradient produced by the imposed temperature gradient. The problem then is to understand the quantity Q* at the atomistic level. There does not exist a theory for the atomistic interpretation of the heat of transport. There are, however, two classes of simple models which seem to afford some physical insight into Q*. Wirtz's model1, z and its subsequent generalizations partition the activation energy, E, for the migration of the atom spatially about the region of the migrating atom, so that Q* is the difference between the portion of the activation energy centered about the original site of the jumping atom and that portion about the arrival site. Clearly, the magnitude of Q* cannot on this model be larger than the activation energy for migration. The model of Oriani3 also considers the spatial distribution of E not only before but also after the atomic jump, and Q* is related to the variation of the spatial distribution of E that is produced by the jump. The simplest kind of system with which to compare the consequences of these ideas is the interstitial solid solution, since one may easily choose a temperature at which only the interstitial component moves, the solvent lattice serving as a completely satisfactory frame of reference. Thus, one avoids complications arising when both species in a binary system move. In addition, one avoids the necessity of knowing the energy for the formation of a vacancy, something which is needed for the analysis of thermal-diffusion data on metals which diffuse by the vacancy-exchange mechanism. However, the number of well-measured interstitial systems is extremely small, and, in particular, there are almost no data in the published literature for the temperature dependence of the heat of transport. If one works with solutes which are gases in the pure state at ordinary temperatures, one may use the technique of thermo-osmosis4 which permits thermal diffusion to be measured over a small temperature difference so that the temperature dependence of Q* can be more easily determined. The choice of dissolved hydrogen in iron and nickel was based on these considerations as well as on the desire to look for an isotope effect associated with localized

Jan 1, 1965

By F. W. Jessen, G. F. Sears

Cavities in massive salt for the purpose of storage of liquid hydrocarbons have assumed a prominent position in recent years. This paper describes a program to facilitate leaching operations for the formation of specifically shaped storage cavities. Various forms and sizes of cavities may be possible through use of the techniques developed. INTRODUCTION The creation of large under ground storage facilities for natural gas and liquified petroleum gases has been practiced for many years. Use of cavities obtained through solution of salt for this purpose is a fairly new and novel approach, but has gained increasing importance due mainly to the economics of this type of mining operation opposed to hard rock mining or surface and pit storage.1 The idea of controlling the shape of any cavity dissolved from massive salt has not been a prime consideration of companies engaged in the formation of storage space. This has been due mainly to an insufficient knowledge of the mechanics of the leaching process and a dearth of published information dealing with both the desirability and ease of control possible for this type of operation. Two distinct advantages are readily ascertainable. From a stability standpoint (i.e., the ability of the completed cavity to withstand stress imposed by overburden pressure and lateral tectonic stresses) a controlled cavity may be generated which will yield the most favorable attitude to these external forces and remain in operable use for longer periods of time.2 A second advantage is that for a given volume, a sphere (which is one controlIed shape possible) represents the minimum surface area exposed. This may become more important in the future when dealing with refrigeration and product losses in underground cavities. The basic solution mining process, without regard for controlling the final shape, is quite simple in that the equipment and materials required to dissolve a cavity in a salt dome or layered salt section are a source of fresh water, a circulating pump, several strings of tubing and a means of disposal of the return brine. To add control measures involves the use of an inert blanket material above the position at which solution proceeds. Initially the annulus of the largest wash pipe is filled with this blanket down to the top of the proposed cavity. The bottom of the proposed cavity is determined by the depth of the original drilled hole or a blanking plug. The blanket material is added incrementally in stages and displaces the water in the enlarging cavity downward. Where the blanket material has displaced the water, no further solution takes place. The rate at which to add this controlling blanket material so as to be able to form specifically shaped cavities of any particular size is of considerable importance. THE PROBLEM It is desirable to have as much information as possible as to the behavior of the mining operation since visual observation obviously is impossible. It would be advantageous, for example, to have a step-by-step program showing how much salt is to be removed, at what rate and the time required. This type of program would facilitate procurement of surface equipment and provide for proper management of manpower requirements, as well as yield a host of intangible benefits. As mentioned earlier, only rule-of-thumb estimates were available until recently and made this type of pre-planned operation haphazardous at best. Recent work3-' has done much to clear up the ambiguities and inaccuracies attendant to solution mining of salt and has helped to place the operation on a more scientific basis. This study attempts to correlate much of the information thus far developed into a complete mining program; to extend the idea of controlled solution mining to include not only spheres, but solid conic sections of the ellipsoidal variety; and to refine the computation of the rate of removal of salt during the various stages of mining. Dommers3 and IIemson4 made extensive laboratory

Jan 1, 1967

By A. K. Snelgrove

This symposium is designed to marshal a state-of-the-art account of approaches and solutions to problems arising out of waterlogging and salinization in the world's largest irrigation system which was dismembered by legal beheading of the Eastern Rivers under the lndus Waters Treaty of 1960. In this introduction, the geological and historical backgrounds of surface waters and groundwaters in the Indus Basin are presented with a summary of remedial measures undertaken and proposed to combat waterlogging and salinization. At a time when water-resource planning is proceeding on a continental scale, a study devoted solely to the Indus River Basin may seem to require an apologia. None is offered, however, for three major reasons: (1) Although the Indus is a single river basin, it is one of great magnitude and importance. The Indus River, which comes from Himalayan sources to form part of the great Indo-Gangetic Plain drainage system, is vital to agriculture and hydroelectric power in the thickly populated northern subcontinent of Indo-Pakistan. (2) The Indus has become a plexus of extraordinary problems since the dismemberment of the world's largest irrigation system of which it was a part. The break-up of this system occurred with the legal beheading of the Eastern Waters under the Indus Waters Treaty of 1960. (3) Th River Basin has been the target of a Herculean effort to contain the twin encroachments of waterlogging and salinization that now threaten it with a shortage of fresh water. The British irrigation system of barrages and canal on the Indus was a venture admirable in conception. Unfortunately, it is now a legacy burdened with unwanted consequences due to interference with the hydrological regimen, particularly the rise of water table occasioned by leakage. To cope with this situation, there has been a possibly unprecedented concentration of domestic and international engineering and scientific talent on the Indus and its problems (about a score of organizations by recent count). To marshal a state-of-the-art account of their approaches and solutions is the purpose of this Symposium. The coordinated effort now being made on a variety of fronts under the general supervision of West Pakistan Water and Power Authority (WAPDA) is directed toward the modification and control of interactions and relationships between geological materials — soils, water and its salt balance, and air —in what has been termed a huge "underground lake" in an alluvial basin. The problem is an intriguing one in that a highly complicated case history of river-course changes visible in the present surface must be projected downwards through Recent and Pleistocene sediments in order to decipher the environmental controls of deposition in terms of vertical and lateral variations in lithology and patterns. Such studies must be extended to compaction, deformation, and fabrics of the alluvium. Geohydrology is often defined as groundwater hydrology. As employed herein, it includes the distribution, quality, and behavior of surface and subsurface waters, including their evolution, vagaries, and geological, climatological, and botanical milieux." GEOGRAPHY The Indus (Sanskrit, Sindhu = river) rises at nearly 17,000 ft above the sea in the glaciers on the northern slopes of Kailas Parbat, Tibet. It is about 1800 miles long and has a total drainage area estimated at 372,000 sq miles. The average discharge is 196,000 cfs at mouth. Leaving the Plateau of Tibet, the river flows at a level which is miles above the sea, and in places miles below Himalayan peaks, in a structurally controlled northwestward course. Joined in this course by six tributaries and swollen by the melt waters of giant glaciers, it swirls about the base of the famous

Jan 1, 1970

By K. T. Koone, R. J. Blackwell

One of the more important problems to be solved in designing a miscible flood is related to the size of the solvent bank used. Size of the bank may be critical to economic success. Too large a bank loses money; too small a bank may deteriorate and fail to maintain the miscibility needed for high recovery. An important factor in deterioration of a small bank is permeability channeling. In a highly stratified reservoir, solvent speeds ahead in the more permeable zones and mixes laterally with fluids bypassed in adjacent, low-permeability strata. Numerical solutions have been obtained .for the differential equations that describe the movement of a slug through a two-layer system in which mixing occurs both in the direction of flow and transversely. The solvent slug is assumed to have the same density and viscosity as the resident fluid and the pushing fluid. These solutions have been verified by comparing them with similar concentration profiles obtained in the laboratory in a 36-ft stratified model packed with glass beads. The theoretical study revealed that when the dominant mechanism causing a bank to fail is lateral mixing the bank size needed for a given recovery may increase with length rather than decreasing as the square root of reservoir length, as suggested by one-dimensional mixing theory. From a comprehensive examination of the variables, a generalized correlation is developed that relates strata thicknesses, bank size, fluid velocity, mixing coefficients, system length and simple solvent-resident fluid phase behavior to the area miscibly swept. INTRODUCTION Miscible displacement, or solvent flooding, continues to receive widespread attention as a method for increasing oil recovery over that possible in conventional gas-drive or waterdrive projects. A basic economic requirement in the application of such processes is the use of as little solvent as possible. A basic physical requirement is that enough solvent be used to maintain miscibility. Economics places an upper limit on the size of a solvent slug, and physical considerations establish a lower limit. Consequently, the practicality of any given miscible process requires that the economic limit be greater than the lower limit imposed by physical requirements. Procedures exist for determining the economic limit; however, procedures for determining realistic minimum bank sizes exist for only special reservoir situations. In the past, bank size has usually been selected on the assumption of a piston-like displacement for which only longitudinal mixing is important. This assumption leads to the favorable conclusion that bank size, expressed as per cent of pore volume, is inversely proportional to the square root of length. Collins considered the problem of transverse mixing of solvent with fluids in bypassed zones with the assumptions that no forward mixing occurs and that the concentration is uniform in the permeable stratum of interest. 1 Lauwerier considered a mathematically similar problem in thermal recovery operations.2 Their work suggests the much less favorable conclusion that bank size could be directly proportional to the length or even higher powers of the length. This paper considers the physical behavior of a small solvent bank as it moves through a real reservoir, without imposing many of the restrictive assumptions of past treatments. To facilitate mathematical description, an idealized, two-layer model that permits mixing both laterally and in the direction of flow will be considered. A calculation procedure for solving the descriptive equations will be developed for a displacement involving fluids of equal density and viscosity. In addition, laboratory experiments designed to check the computational results will be described, and the coincidence of calculated and observed results will be discussed. The use of many solutions for a variety of bank sizes, strata thicknesses and characteristic system lengths to develop a usable correlation between

Jan 1, 1966

By C. W. Volek, J. E. Chappelear

The injection of a hot liquid into initially cool porous media, saturated with the same liquid and surrounded by two impermeable but heat conducting media (cap and base rock), has been studied both experimentally and theoretically. The temperature dependence of the viscosity was included in the theoretical model, but it was assumed that the specilic heats and densities of the various materials were independent of the temperature. Solutions to the theoretical model were approximated by numerical methods. Both theoretical and experimental results indicate that center-line temperatures are significantly higher than boundary temperatures. Comparison of experimental and theoretical results with a cold/hot viscosity ratio 01 19:1 were in reasonable agreement. Theoretical calculations show that the effect of the temperature dependence of viscosity was very significant at ratios of 100:l to 1000:1, which are typical of those that occur when injecting hot water to flood heavy oil reservoirs. INTRODUCTION We consider the problem of prediction of fluid flow and temperature distribution in an initially cold-fluid-filled reservoir on the injection of the same hot liquid by the use of mathematical and physical models. The results reported are for a two-dimensional rectangular section of the reservoir, as shown in Fig. 1. The injection and withdrawal faces are assumed to be equipotentials for fluid flow. The ultimate purpose of such models would be to predict hot-water injection performance. However, we note that in the work presented here, one of the most significant aspects of the problem — the instabilities resulting from two-phase, water and oil flow — is not included. We will not give a historical review, but refer instead to the paper of Spillette and Nielsen,1 which contains a rather complete bibliography and critical discussion. The physical problem is exactly the same as Spillette and Nielsen, except for certain simplifications in our assumptions. The mathematical details are somewhat different, and we will present the details of our method here. The mass flow equation, which is elliptic in character, is handled by successive overrelaxation. The heat flow equation, which is parabolic in character, is handled by a straightforward explicit approximation. Some difficulty arose in the over-all heat balance due to small errors in the solution of the mass flow equation, and we feel that a different formulation of the heat flow equation would be desirable for future work. In addition to the mathematical solutions for the temperature distributions in a porous medium due to the injection of hot liquids, experimental data are presented to check the validity of these solutions. A schematic diagram of the model is shown in Fig. 2. Certain qualitative physical conclusions were obtained from our numerical and experimental results. These are: 1. Assuming high (infinite) conductivity normal to the bedding plane in the reservoir is a poor approximation, and may lead to overestimates of the total heat losses (to cap and base rock) of as much as 50 percent. 2. More heat is retained in the reservoir (per unit of heat injected) for higher viscosity changes. 3. Any particular temperature isotherm moves more rapidly along the center line for higher viscosity changes. Consequently, the approximation of temperature independent viscosity is not suitable for obtaining quantitatively correct results. MATHEMATICAL MODEL Our mathematical model is that of a reservoir of thickness 2h. The problem is idealized from that of a linear hot-water drive, and we imagine that the input face is sufficiently far away from the injection wells that the stream lines enter it normally. Similarly the production wells are far enough from the outflow face that the flow lines leave it normally. The reservoir and surrounding cap and base rock

Jan 1, 1970

By H. J. Morel-Seytoux

Methods of predicting the influence of pattern geometry and mobility ratio on water flooding recovery predictions are discussed. Two methods of calculation are used separately or concurrently. The analytical method yields exact solutions in a convenient form for a unit mobility ratio piston-like displacement. A few typical pressure distributions, sweep efficiencies and oil recoveries are presented for various patterns. For non-unit mobility ratio, one may resort to a numerical method, such as that of Sheldon and Douherty. 1,2 Because the domains of applicability of the analytical and numerical techniques overlap, the exact solutions provide estimates of the errors in the numerical procedures. The advantages of the analytical and numerical methods can be combined. To develop a numerical technique as independent of geometry as possible, the physical space is transformed into a standard rectangle. The entire effect of geometry is rendered through one term, the "scale-factor", derived from mapping relations. The scale factor can be calcu-lated from the exact unit-mobility ratio solution for the particular pattern of interest. By this means recovery performances for arbitrary mobility ratio con be obtained for many patterns. A sample of tesults obtained in this manner is presented. INTRODUCTION Pattern geometry and mobility ratio are two major factors in making a waterflood recovery prediction. Because assisted recovery has become increasingly important to the oil industry, pattern configuration and mobility ratio also assume a greater significance in the assessment of the economic value of recovery projects. The influence of pattern geometry and mobility ratio in shaping a recovery curve and on the other quantities of interest to the reservoir engineer is the main subject of this paper. Much effort has already been spent on estimating quantitatively the influence of either pattern or mobility ratio or both on oil recovery. The literature reports many investigations of this nature. 3-9 However, many results or methods of recovery prediction presented in the literature cannot be considered fully satisfactory. Even for unit mobility ratio and piston-like displacement, where analytical solutions are available, the literature shows discrepancies. For non-unit mobility ratio, the divergence in the results is extreme. For infinite mobility ratio in a repeated five-spot, depending on the investigator, the sweep efficiency ranges from 0 per cent to 60 per cent. With respect to the influence of pattern on recovery, only the repeated five-spot has received much attention. Other confined patterns and pilot configurations have received very little attention. Two calculation methods are presented in this paper, either separately or concurrently: the analytical method of potential theory and the numerical method of finite-difference approximation. The analytical method is more restricted in scope than the finite-difference method, but it has the definite advantage of providing exact solutions within its range of applicability. If a unit-mobility ratio piston-like displacement is assumed, the analytical approach is possible. A few typical results are reported in this paper; the detailed description of the general method and of a great variety of results will be the subject of other articles. Fornon-unity mobility ratio, we must resort to a numerical scheme. The numerical technique is that which was described by Sheldon and Dougherty.l,2 It is not limited to piston-like displacement. However, mainly single interface results will be presented here. Because the respective domains of applicability of the analytical and the numerical method overlap, useful comparisons of exact and numerical solutions can be made for a variety of patterns. The advantages of the analytical and numerical approaches can be combined. The reason for the success of this analytical-numerical approach can be summarized in the following two points: 1. The numerical solution for arbitrary mobility ratio can be programmed most efficiently when the physical space in which the displacement actually takes place is transformed into a standard shape; and 2. This can be done with remarkable simplicity whenever an analytical expression for the pressure

Jan 1, 1966

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Ranaome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 3Y4x41/4-in. tool joints; 8-in. bits, 23/4x 3 3/4-in. tool joints; 6-in. bits, 21/4x31/4-in. tool joints; 4-in. bits, 15/ix25/s-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 2-in. drill pipe with tool joints, 31/2-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 5 to 37/8 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Rannome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 31/4x41/4-in. tool joints; 8-in. bits, 23/4x 33/4-in. tool joints; 6-in. bits, 2Y4x3Y4-in. tool joints; 4-in. bits, 15/ix25/8-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 23/8-in. drill pipe with tool joints, 31h-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 55/8 to 3 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a 'moral victory' as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp's brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953

By R. F. Mehl, C. E. Birchenall

R. E. Hoffman and D. Turnbull—The authors have presented evidence which they have interpreted as indicating that the rate of self diffusion is not intrinsically more rapid at grain boundaries than within the grain. Grain-size effects which apparently exist are attributed rather to impurities concentrated at the grain boundaries. In view of our own experiments and the existing evidence, we believe that the support for this hypothesis is not convincing. We have in progress an investigation in which the rate of self diffusion of silver is being measured over an extended temperature range in both single-crystal and polycrystalline specimens. The results of the single-crystal experiments and some preliminary data on fine-grained polycrystalline specimens have already been reported:' and it is anticipated that a complete report will be published in the near future. The self-diffusion coefficient of large-grained polycrystalline silver (1 grain per sq mm) has previously been measured by Johnson" between 730" and 940°C. The diffusion coefficients which we have measured in single crystals (210 plane normal to diffusion direction) agree within experimental error with values calculated from an extrapolation of Johnson's curve down to temperatures as low as 500 °C. However, it has been demonstrated that the overall self-diffusion rate in fine-grained polycrystalline specimens (initial grain size of 0.003 cm) becomes measurably larger than the overall rate in a single crystal at a temperature of 600°C, and the discrepancy between the two rates becomes greater as the temperature is further decreased. In fact, it has been possible to obtain satisfactory penetration curves for polycrystalline specimens using the sectioning technique at temperatures as low as 400°C. At this temperature, the penetration is 50 to 100 times greater in the polycrystalline specimens than in a single crystal. Fisher" has developed an analysis whereby the ratio of the rate of the unit diffusion process at the grain boundary to the corresponding rate within the grain can be calculated from the penetration curves and an assumption as to the width of a grain boundary. This analysis applied to our data indicates that the unit process at the grain boundary is faster by a factor of 10' at 475°C when the grain boundary width is taken to be 5. The silver used in most of these experiments was obtained from the Handy and Harmon Co. and listed as 99.97 pct pure. Preliminary experiments on 99.999 pct silver from the Jarrel-Asch Co. indicate a grain-size effect of the same order of magnitude as in the less pure silver. Nominally, these purities are as good, at least, as that of the carbonyl iron used by the authors, but of course if an impurity effect does exist its magnitude might be very dependent upon the nature of both major and minor constituents. The authors have cited the work of other investigators who have found no grain-size effects. Neither Steigman, Shockley and Nix" nor Maier and Nelson8 were able to correlate self-diffusion coefficients of copper with grain size. However, all their measurements were performed at or above 750°C; and on the basis of our work with silver, no grain-size effect would be expected at temperatures above about 0.7 of the absolute melting temperature unless the grain size were exceedingly small. Likewise, in the investigation of the self diffusion of lead by Seith and Keil,14 the lowest temperature at which the diffusion coefficient was measured in polycrystalline specimens was 207°C, which is still sufficiently high so that the lack of a grain-size effect is not surprising. Finally, in those experiments on iron from which they concluded that there was no grain size effect, Drs. Birchenall and Mehl seem to have no information as to the actual grain sizes immediately prior to and following the diffusion anneal. Without this information, we believe that their own experiments offer little support for their hypothesis. F. S. Buffington, I. D. Bakalar, and M. Cohen—The results given in this paper agree in order of magnitude with those tentatively reported by us.27 However, significant differences exist in the two sets of data, and it may be well to make an explicit comparison. The diffusion studies at M.I.T. were conducted on somewhat higher purity iron (99.98 pct Fe) than the grades used by the authors, but this is undoubtedly not the answer. Fig. 4 shows the diffusion results of both laboratories for the gamma phase, omitting the authors' data on the commercial steels, while fig. 5 presents a similar comparison for the alpha phase. The divergence is much more marked in the latter case than in the former. In connection with the M.I.T. determinations, all of the runs in the gamma range and those above 800 °C in the alpha range were conducted with specimens having a relatively thick (0.002 cm) electrodeposit of radioactive iron. This practice minimizes any possible error due to extraneous diffusion that may occur during the heating to and cooling from the operating temperature. An exact solution of Fick's law for these boundary conditions was used in calculating the diffusion coefficients. At a later time, three runs were made below 800°C, using very thin electrodeposits similar to those of the authors, and the points fell considerably below the values expected from the extrapolation of the results based on the specimens with the thick deposits (compare dash-dot line in fig. 5). However, in the runs with the thin deposits, deviations of 100 pct were found between the individual specimens, whereas the maximum deviation with the thick deposits was less than 25 pct. Accordingly, it is not known at the moment whether the M.I.T. points below 800 °C should be given as much weight as those above 800°C. If this were done, the frequency factor would be of the order of 400 cm2 per sec, which is quite high. In other metals, the frequency factor for self diffusion lies between about 0.1 and 10 cm2 per sec. As the points below

Jan 1, 1951

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Ranaome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 3Y4x41/4-in. tool joints; 8-in. bits, 23/4x 3 3/4-in. tool joints; 6-in. bits, 21/4x31/4-in. tool joints; 4-in. bits, 15/ix25/s-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 2-in. drill pipe with tool joints, 31/2-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 5 to 37/8 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a 'moral victory' as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp's brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Rannome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 31/4x41/4-in. tool joints; 8-in. bits, 23/4x 33/4-in. tool joints; 6-in. bits, 2Y4x3Y4-in. tool joints; 4-in. bits, 15/ix25/8-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 23/8-in. drill pipe with tool joints, 31h-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 55/8 to 3 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By H. I. Smith

WHEN Spain established colonies on the North American continent, some of her land grants, in what is now the United States, reserved to the Crown deposits of gold, silver, and mercury. Later mineral rights were reserved under some of the English Crown charters involving land in the eastern part of North America, with provisions for the payment of royalties thereon to the Crown. Reservation of mineral deposits to the United States Government was instituted by an ordinance of Congress on May 20, 1785, which applied to such deposits in the Northwest Territory, then north of the Ohio River and east of the Mississippi River, and provided that there should be reserved "one third of all gold, silver, lead, and copper mines, to be sold or otherwise disposed of as Congress shall hereafter direct." Little was known then of the mineral resources of the country; the Great Lakes copper region had just come into the possession of the United States by treaty and much of the western mineral land still belonged to France and Spain. The policy of leasing mineral deposits was enacted by Congress under the act of March 3, 1807, which provided that "the President of the United States shall be, and is hereby, authorized to lease any lead mine which has been or may hereafter be discovered in the Indiana Territory, for a term not exceeding five years," and in the same year the Government reserved 345,600 acres of land in northern Illinois, valuable for lead. In 1816 Congress provided that in all cases where a tract of public land containing a lead mine or salt spring was applied for by settlers on the public domain, no permission to work the mine or spring would be granted without the approbation of the President of the United States. By the act of March 1, 1847 (9 Stat. 146), the control of mineral lands was transferred, with all records, from the War Department to the Treasury Department, and by the act of March 3, 1849 (9 Stat. 395), supervisory powers over lead and other mines of the United States were transferred to the Secretary of the Interior. Following the discovery of gold in California in 1848, President Polk advocated the leasing of mineral lands acquired from Mexico under the treaty of 1848. However, owing to the lack of communication and transportation facilities and the consequent difficulty of checking on production and operations, leasing was found impracticable. Mining was permitted to be carried on in each district under rules made by the miners themselves and patterned after the Mexican mining system. To meet the situation, Congress enacted mining laws in 1866, 1870, and 1872, which provided for possession by location and for private ownership, after discovery, by patent of essentially all mineral deposits belonging to the United States, except coal. Many other acts were passed by Congress applying to specific regions and states, reserving salt, lead, or other minerals, some of which provided for the leasing of such deposits. However, very few if any leases were issued thereunder. Not until 1901 (31 Stat. 745) were the mining laws extended to include salt, and in 1910 they were revised to specify salines and associated products. At the turn of the century, it became more and more evident that the mining laws developed for metallic minerals were not practical for the development and conservation of oil and gas, coal, potash, phosphate, or oil shale. Following withdrawals by the Secretary of the Interior as authorized by the act of June 25, 1910, the Department of the Interior recommended leasing legislation; however, it was not until February 25, 1920, that the Mineral Leasing Act was passed by Congress. This act authorized leasing all public lands potentially valuable for oil, gas, coal, phosphate, sodium, or oil shale. The Organic Act, creating the Geological Survey in 1879, imposed upon its director the duty of classifying the public lands.' The early years of the Survey were devoted largely to the accumulation of fundamental data and, with only minor exceptions, land classification was not seriously undertaken until 1906. Since that year, it has been actively pursued with respect particularly to leasable minerals and water power values. On March 3, 1873 (17 Stat. 607) Congress authorized the sale of coal lands of limited acreage to individuals at $10 to $20 an acre, depending on the distance from a railroad, and in 1907 a new scale of prices was adopted by the Secretary of the Interior, based, more logically, on the quality and thickness of the coal deposits, their depth below the earth's surface, and their accessibility. At that time it was much cheaper and easier to obtain incidental title to coal rights under the homestead laws at $1.25 an acre. On March 3, 1909, the first separation act (35 Stat. 844) became a law. It authorized patents, with a reservation of the coal and mining rights involved to the United States, to persons who in good faith had entered public lands under the non-mineral land laws prior to withdrawal, classification, claim, or report that such lands were valuable for coal. It solved the problem only partially, and the practice thereupon adopted by the department of making coal withdrawals "from all forms of entry" instead of "from coal entry" expedited the more complete solution effected by later acts. By the act of June 22, 1910 (36.Stat. 583) withdrawn and classified coal lands were declared subject to entry under the homestead, desert land, and Carey acts, provided a waiver of the coal rights accompanied the application, and by the act of April 30, 1912 (37 Stat. 105) the same privilege was extended to State selections and isolated tracts. The act of August 24, 1912 (37 Stat. 496) extended the separation policy to lands withdrawn for oil and gas in Utah, and finally the act of July 17, 1914 (38 Stat. 509) extended it to all nonmineral filings on public lands theretofore or thereafter

Jan 1, 1955

By Paul A. Beck

THE occurrence of the (100) [lOO] or "cube" texture upon annealing of cold-rolled copper has been much investigated.' The conditions favorable for its formation were found to be a high final annealing temperaturez or long annealing time," a high reduction of area in cold rolling prior to the final anneal,' and a small penultimate grain size." The effects of penultimate grain size and of rolling reduction were found by Cook and Richards4 to be interrelated in such a way that any combination of them giving lower than a certain value of the final average thickness of the grains in the rolled material leads to a fairly complete cube texture with a given final annealing time and temperature. Also, according to the same authors, at a higher final annealing temperature a larger average rolled grain thickness, i.e., a lower final rolling reduction, is sufficient than at a lower temperature. These somewhat involved conditions can be understood readily on the basis of recent results obtained at this laboratory. Hsun Hu was able to show recently by means of quantitative pole figure determinations that the rolling texture of tough pitch copper, which is almost identical with that of 2s aluminum: may be described roughly as a scatter around four symmetrical "ideal" orientations not very far from (123) [112]. In the case of aluminum, annealing leads to retain-ment of the rolling texture with some decrease of the scatter around the four "ideal" orientations, and to the appearance of a new texture component, namely the cube texture." A microscopic technique, revealing grain orientations by means of oxide film and polarized light, showed that the retainment of the rolling texture is achieved through two different mechanisms operating simultaneously, namely "re-crystallization in situ," and the formation of strain-free grains in orientations different from their local surroundings, but identical with that of another component of the rolling texture. Thus, a local area in the rolled material, having approximately the orientation of one of the four "ideal" components of the texture, partly retains its orientation during annealing, while recovering from its cold-worked condition, and it is partially absorbed at the same time by invading strain-free grains of an orientation approximately corresponding to that of another "ideal" texture component. The reorientation here, as well as in the formation of the strain-free grains of "cube" orientation, may be described as a [Ill] rotation of about 40°, see Fig. 1 of ref. 6. The preferential growth of grains in such orientations is a result of the high mobility of grain boundaries corresponding to this relative orientation.' " It appears very likely that in copper the mechanism of the structural changes during annealing is similar to that observed in aluminum (except for the much greater frequency of formation of annealing twins in copper). In both metals the new grains of cube orientation have a great advantage over the new grains with orientations close to one of the four components of the rolling texture. This advantage stems from their symmetrical orientation with respect to all four retained rolling texture components of the matrix; they are oriented favorably for growth at the expense of all of these four orientations. As a result, the growth of the "cube grains" is favored over the growth of the others, as soon as the new grains have grown large enough to be in contact with portions of the matrix containing elements of more than one, and preferably of all four component textures. It is clear that this critical size is smaller and, therefore, attained earlier in the annealing process if the structural units, such as grains and kink bands, representing the four matrix orientations are smaller, i. e., if the average thickness of the rolled grains is smaller. Hence, for a given annealing time and temperature, a smaller penultimate grain size and a higher rolling reduction both tend to increase that fraction of the annealing period during which the above condition is satisfied. Consequently, the percentage volume of material assuming the cube orientation increases. The same is true also for increasing time and temperature of annealing when the penultimate grain size and the final rolling reduction are constant, since the average size attained by the new grains during annealing increases with the annealing time and temperature. For the same reason, at higher annealing temperatures a given volume percentage of cube texture can be obtained with larger rolled grain thickness (larger penultimate grain size, or smaller rolling reduction) than at lower annealing temperatures. The well-known conspicuous sharpness of the cube texture may be interpreted as a result of the fact that selective growth of only those grains is favored that have an orientation closely symmetrical with respect to all four components of the deformation texture and exhibit, therefore, a high boundary mobility in contact with each. The effect of alloying elements in suppressing the cube texture, as described by Dahl and Pawlek,' appears to be associated with a change in the rolling texture. For face-centered cubic metals, such as copper, which do exhibit the cube texture upon annealing, the rolling texture is always of the type described above, i. e., scattered around four "ideal orientations" of approximately (123) [112]. The addition of certain alloying elements, such as about 5 pct Zn or 0.05 pct P in copper, has the as yet unexplained effect of changing the rolling texture into the (110) 11121 type. This texture consists of two fairly sharply developed, twin related components. In such cases, as in 70-30 brass and in silver, the annealing texture again is related to the rolling texture by a [lll] rotation of about 30°, however, because of the different rolling texture to start from, it has no cube texture component. At higher temperatures, both in brassm and in silver," grain growth leads to a further change in texture: A [lll] rotation of the same amount, but in reversed direction, back to the original rolling texture.

Jan 1, 1952

By Peter R. Drummond

THE final casting of refined copper has been re-J- stricted for generations by the following sequence of operations: Filling the reverberatory furnace, melting, skimming, blowing or flapping, and poling. The hoped-for 24 hr cycle, producing 300 tons or more, has been taken up largely with the necessary bat time-consuming tasks of cleaning the bath, sulphur elimination, and in turn removal of excess oxygen to produce tough-pitch copper. Incidental to comparatively slow melting under combustion gases, copper oxides react with the furnace lining, and the slag so-formed must be completely recycled. The three-phase arc furnace has eliminated some of the cycle stages, and telescoped the remainder into a continuous operation. Electrical energy, supplied to graphite electrodes enclosed in high grade refractories, rapidly melts copper cathodes and sustains a stream of metal, containing approximately 0.01 pct oxygen, without contamination from fuel. The arc was struck on the first large electric furnace for melting copper in the United States on April 13, 1949. The earliest use of this type of furnace was at Copper Cliff, Ont., in 1936, and an admirable description of their installation has been published? Copper, melted in the Baltimore furnace, is used to cast billets, and the installation differs somewhat from the Canadian, as will be described. The arc furnace is a heavy-duty, three-phase furnace, holding 50 tons, the general outline of which appears on Fig. 1. The steel shell is 15 ft ID with a bottom radius of 14 ft 2 in. The roof, separate and distinct from the body, consists of a 15-ft water-cooled, cast-steel ring of the same outside diameter as the furnace. The center line of the furnace lies 9 ft 6 in. from that of the trunnions, permitting a 5" backward tilt for skimming, and a 40" maximum nose tilt forward for complete draining. Normally, the furnace overflows by displacement, and the use of the forward tilt arrangement is restricted to covering charging delays. The charging slot, 3 ft 8 in. x 5 in., lies on the north center line, the tap hole on the south, and the 30x30 in. skim door 45" to the west of the slot. The original 20-in. graphite electrodes were replaced with 14 in. in December 1949. Three conventional direct current winch drives, governed by electrical controls, position each electrode which has individual mast supports and counterweights. An independent circulation supplies cooling water for the electrode glands, the roof ring, charge slot, and the skim door frame. Arc Furnace Refractories Hearth: Fused-in monolithic bottoms had been used in copper arc furnaces, installed prior to April 1949. These consisted of thin layers of periclase, successively fused in place over preliminary brick courses. Heat was obtained from the arc, using a T-like arrangement of broken electrodes resting directly on the periclase to be fused. The operation, taking weeks to perform, was very expensive. The chemically-bonded magnesite-brick bottom, installed at Baltimore, was the first of its kind and a radical departure from previous practice. It consists of a 1 to 6-in. layer of castable refractory laid on the steel shell, modifying it to a 12 ft 2 in. bottom radius. Two courses of 9x2 % -in. fireclay straights and keys follow. The third course is made of 9-in. magnesite blocks of special shape to form circles of an inverted arch. It was started by a four piece keystone with skew-backs forming the outer course. The fourth course also started on a central keystone, or button, of four 90" segments, 12 in. diam x 13 Vz in. deep, and continued with 13%-in. blocks. Skewbacks at the shell completed the course to produce a horizontal surface for the side walls with a single course of No. 2 arch fireclay against the steel. Dry chrome-magnesite cement was brushed over each course after laying, and a 1-in. expansion space between the brick and the shell was filled with the same mixture. The total bottom thickness, excluding the castable material, was 5 in. of clay plus 22% in. of chemically-bonded magnesite. Tap Hole: A 5-in. OD and 3-in. ID silicon-carbide tube constitutes the tap hole and is set tangential to the upper course of the furnace bottom. Molten metal fills the tube when the furnace is level and filled to capacity. Side Walls: The lining, against the shell, consists of a 9x4Y2x3 in. soldier course of fireclay, using straights and No. 1 arches to turn the circles. A second soldier course of 9x4'/2x2'/2-in. fireclay was laid in a somewhat similar fashion. Three courses of 13Y2x6x3 in. and 9x6~3 in. of final magnesite, laid flat, completed the lining, using Nos. 1 and 2 keys to turn the circles. Cardboard spacers were placed between every two bricks in horizontal courses, and a thin coat of chrome-magnesite cement filled the joints between the firebrick and magnesite. A sprung-arch spanned the skim door with jambs of suitable magnesite shapes. Charge Slot: The slot is 3 ft 8 in. wide x 5 in. high. A silicon-carbide sill of special shapes has a 30" slope to allow cathodes to slide easily into the bath. The original arch was flat, and composed of Nos. 1 and 2 wedge magnesite with a 6-ft radius. It projected 5 in. over the sill, and, being a flat arch, gave an 18 15/16-in. opening between the inner edge and the metal line. The whole assembly was later raised 9 in., and the flat arch replaced with an arch, the lower edge of which maintained the 5-in, width from the outer to inner edges as shown in Fig. 2. A water-cooled, cast-copper jacket protects the steel shell behind the slot.

Jan 1, 1952

By A. L. Bailey, B. A. Landry

W. W. ANDERSON and G. E. KELLER*—We want to compliment the authors on this very thorough paper. It gives information which the coal industry has needed for some time. We hope that the additional information which the authors are collecting will he available shortly. The mixing and riffling procedure that was followed for experimental purposes is obviously not practical in routine float-and-sink testing because of the particle size degradation which would result in handling the sample so many times. It is important to obtain our tloat-and-sink fractions with a minimum amount of handling of material. A statement is made in the paper (p. 80) that "the variable most likely to affect the size of sample required to meet a given preassigned accuracy would be the state or degree of mixing of the coal." We agree that this is a large factor, but do not believe it is the most important factor. Our own opinion is that the most important single factor governing the total gross weight of sample that must be collected is the percentage of the weight of material in the smallest fraction that results from the screening and float-and-\ink operations. In other words, size of sample is governed by the total number of fractionations that must he made, and the distribution of material within the fractions. We can imagine a coal with perfect mixing, but with such a small amount of material in some float-and-sink fraction in one of the coarse sizes that a much larger sample would have to be taken than would be the case with very poorly mixed material, but with a large percentage of coarse material more evenly distributed in all float-and-sink fractions. Our own observation of many float-and-sink tests that we have run in our own organization on many types of coal is that the size of sample that must be used on fine size float and sink is governed more by the requirements for weight of material to be used for analysis in the laboratory than by weight of material necessary to obtain accurate float and sink percentage of weight values. In other words, it is our opinion that very small samples can be used for float-and-sink fractionation in the fine sizes, but that accurate analysis of the fractions will depend on a larger weight of sample being pulverized for the laboratory than is necessary to establish the float-and-sink distribution with respect to weight. A. L. BAILEY and B. A. LANDRY (authors' reply)—The authors thank Messrs. Anderson and Keller for their comments based on long experience. It is agreed that the involved mixing and riming technique used may be disadvantageous from the standpoint of degradation. Fortunately, the paper does point out that the extended riming was unrewarding in causing further mixing. Two large unknowns remain, however: (1) how much of the mixing from the presumed highly unmixed state in the bed was achieved toward the random state during blasting, loading, transportation, screening, and further transportation to the point where the gross sample was taken, and (2) how much of the mixing took place during the preparation described preceding riming. As has been pointed out by one of the authors.6 the degree of mixing has a very large effect on the size of sample required and there are still too few experimental data to show at what stage of coal handling most of the mixing occurs. The discussion states that the weight of material in a screened fraction, or in a float-and-sink fraction, is more important than the mixing factor. We do not believe that these factors are comparable in this instance inasmuch as our purpose was to give minimum sampling requirements to achieve a preassigned accuracy in the percentages of float, middlings, or sink, and nothing more. The gross sample had already been screened and no further division by screening was made or contemplated; also, it was not intended that the middlings and sink fractions would necessarily be adequate for percentage ash or other determination. In other words, the sample obtained by the method outlined is not intended for washability studies but only for preparation plant control. Further experimental work has been done, since the paper was prepared, to investigate the effect of increasingly larger top and bottom sizes on the variability of float, etc., of a double-screened coal from Western Pennsylvania. Results will be published and eventually attention is to be given to the preparation of sampling specifications. E. H. M. BADGER*—I should like the authors to explain more fully the fundamental assumptions on which their Eq 4 is based. The equation is of the form s2 = p(l - p) which is the usual expression for the (standard deviation)2 when the chance of finding a particular kind of particle in the sample is proportional to the number fraetion, p. But instead of the number fraction, the authors have used the weight fraction, WF/W. The chance of finding a particular kind of particle in the sample can only be proportional to the weight fraction, if the average ?eig?ts of all kinds of particles, that is, float, midlings, or sink, are the same. Surely a much more justifiable assumption would be that the average volumes of the particles are the same, and, if this is so, Eq 4 would not be true. This may be demonstrated as follows: Let be the weight fraction of float, middlings, or sink, dl the density of this fraction, and d2 the density of the rest of the coal. Then assuming that the average volumes of the pieces in the three classes are the same, the number fraction, p, is given by ? P = d1/l-?/d2 + ?/d1 = ?d2/d1 + ?(d2-d1) The weight fraction, w, in terms of p is given by ? = pd1/(l-p)d2 + pd1 = pd1/d2 + p(d1-d2) _____ [61

Jan 1, 1950

By A. H. Geisler, F. Keller

Although the subject of precipitation from solid solution appears to be one of the more profitable fields in metallurgy for study with the electron Microscope, few comprehensive studies have yet been made. Occasionally electron micrographs have been published that illustrated alleged precipitation in various alloys, but frequently the apparent size, shape and distribution of the particles do not fully agree with those expected from theory or from observations of the Precipitate when the particles have grown to a size resolvable with the light microscope. The presence of a Widmanstatten pattern for submicroscopic precipitate particles has been demonstrated for only a few aged a1loys.l The initial problem has been the development of suitable techniques for applying the transmission-type instrument to a field that has been inherently associated with reflection-type microscopes. The various techniques have been the subjects of numerous publications, and instead of describing them individually here it will sufficc to say that the oxide film method has proved to be the most satisfactory method yet found for studying the microstructure of aluminum alloys with the electron microscope. This method has the definite characteristic advantage that the actual surface layer of the specimen is examined and not a plastic or silica replica. Doubtless suitable methods for forming and removing the oxide film could be developed for alloys of other metals. The purpose of this report is to point out the characteristics of the oxide film method that have been observed during the studies of numerous aged aluminum alloys and to present the results of these studies. Preparation OF Specimens The oxide film method for preparing specimens of aluminum alloys for examination in the electron microscope has been described in detail previously,¹,² Briefly, the procedure consists of preparing the metallographic specimen, forming the film by anodic oxidation and removing the film from the prepared surface of the oxidized specimen, The specimen is polished accordillg to the usual procedure for microscopic examination.³ he specimen may then be etched to remove flowed metal but etching to attack the alloy constituents or to leave them in relief as in the replica processes is not necessary. Electrolytic polishing is not generally recommended, Deep-etched specimens are used frequently, however, since they provide information that is not revealed by polished specimens and frequently present the same information more clearly than do the polished specimens, The surface preparation of specimens to be deep-etched is not important, since specimens that have been subjected to the first wet PoliShing operation are generally used, but frequently the as-rolled surface is Suitable. These specimens are then etched using a suitable macroetching reagent.

Jan 1, 1947

By E. P. Klier, Taylor Lyman

THE structures formed when austenite is quenched to subcritical temperatures and allowed to transform isothermally have been the subject of intensive study since the work of Davenport and Bain.1 Isothermal transformation diagrams summarizing the results of many of these studies have become widely familiar to metallurgists. Of particular interest to metallographers are the dark-etching, acicular products of transformation formed by isothermal reaction below the temperature of maximum velocity of the austenite to pearlite reaction. These products, the bainite structures, can be formed in a wide variety of steels. However, the constitution of bainite is not well understood and divergent views have been expressed as to its mode of formation. In this investigation a combination of dilatometric, microscopic and X-ray methods has been brought to bear upon the problem in the hope of some elucidation of the bainite reaction as it occurs in hypoeutectoid steels. MECHANISM OF BAINITE FORMATION Davenport and Bain1 considered the mechanism of bainite formation to consist of two steps-an allotropic transformation followed by precipitation of carbide. The separation in time of the two processes was considered to be very slight in the upper temperature range and very marked at temperatures just above the martensite range. This concept has been restated by Vilella, Guellich and Bain2 in the form of a definition of the acicular mode of transformation as: The successive, abrupt formation of flat plates of supersaturated ferrite along certain crystallographic planes of the austenite grains; this supersaturated ferrite begins at once to reject carbide particles, (not lamellae), at a rate depending upon temperature. In effect, this is the acicular mode of transformation, even though the temperature be such as to limit the actual life of the quasi-martensite to millionths of a second. The investigation of isothermal decomposition of austenite in certain alloy steels (notably those containing chromium or molybdenum) has revealed that there may be a range of temperatures between the pearlite and bainite reactions in which the austenite decomposes at a relatively low rate.3-6 Further, it is characteristic that in the second region of fast reaction the decomposition of the austenite is incomplete by one reaction but goes to completion by a second reaction. These observations led Wever7 to a description of the bainite reaction as follows: I. The reaction takes place by initial precipitation of a martensitic intermediate structure. 2. In the upper temperature range this structure readily decomposes into ferrite and cementite. In the lower range carbon separates in an unknown form. 3. In the upper temperature range the precipitated carbide nucleates the precipitation

Jan 1, 1944