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By David L. Sikarskie

The measurement of residual stress fields using X-ray diffraction techniques requires the removal of successive layers of material from the specimen. In the process of removing these layers, the underlying residual stress field is altered; hence, if the original stress field is desired, corrections must be made to the measured stresses. Although these corrections already exist for a number of cases, an alternate procedure is outlined herein whereby the corrections are presented in series form. This form is numerically more convenient within certain ranges of correction depths and in its simplicity gives additional insight into the factors affecting the corrections. It, further, readily permits determining the proper slice depth to remove and the effect on the correction of both specimen thickness and magnitude and form of the measured residual stress itself. ONE of the techniques currently being used to determine residual stress fields is X-ray diffraction. Since X-rays are diffracted essentially from the layers of atoms near the surface of the specimen, of the order of 0.0001 to 0.001 in. depending on conditions, the method can only determine stresses* in this very thin surface region. Therefore, if the residual stress field at some depth into the specimen is required it is necessary to remove successive layers of material taking measurements after each cut. In the process of removing a thin surface layer of stressed material, however, the over-all equilibrium of a cross section of the specimen is disturbed with the result that the internal residual stress which existed before the slice of material was removed is altered. This alteration of stress occurs after each slice and its effect is cumulative. Thus, if the original stress distribution is desired, corrections must be applied to the measured stresses. These corrections can be found by satisfying equilibrium on each new cross section. These corrections have been derived in detail for a number of cases, e.g., cylinders, bars, plates,' and also are given in Appendix VI of Ref. 2. In each of the above cases the correction terms require the evaluation of definite integrals of the measured residual stress distribution. In general, this is indeed the way one would proceed with the corrections. There are a number of instances, however, where the integrals can be replaced by a series form of correction which is considerably more convenient numerically and, also, gives added insight into the factors which influence the corrections. Such a form of the correction is particularly useful, for example, if one is interested in designing a test such that stress corrections will be small. Also, the question of what depth of slice to use in relation to both specimen thickness and residual stress distribution itself must invariably be answered. Questions of this type can be readily answered using a series form of correction. In addition, if the correct residual stress distribution is needed only for limited depths into the specimen the series correction provides numerical results which are equivalent to the integral correction. I) DERIVATION OF THE SERIES CORRECTION The problem to be considered here is that of a bar or plate,* i.e., either one or two dimensions of the specimen is large compared with the third. Also, the residual stress distributions considered depend only on z, see Fig. 1. In view of this it can be shown' that a stress correction, based on simple beam theory, satisfies all conditions of "elasticity theory" and hence is the correct (and only) solution. The integral form of the correction (this assumes that the measured residual stress is a piecewise continuous function) as given in Ref. 1 is:

Jan 1, 1968

By Robert G. Nisle

In conducting a pressure build-up test on an oil well, it is often necessary to shut-in the well for a short time prior to initiation of the test. The effect of such a short tern shut-in on the resulting pressure build-up curve is derived. The derivation is based on the Kelvin point-source solution and is a straight forward extension of the fundarnental theory. The magnitude of the effect is evaluatecl and illustrated by an example. It is shown that if the production subsequent to the short term shut-in is at least 10 times the duration of the shrct-in, the effect will he less than 10 per cent. INTRODUCTION It often happens that before a pressure build-up test can be made an additional valve must be mounted on the wellhead, or on arrival at the well, it might be learned that the well had been shut-in several hours the previous day because of a shortage of tankage. In such cases the question is How will such an interruption of production affect the results of a pressure buildup test This work explains how to take account of such a short term shut-in in order that the test can proceed without further delay. The solution presented here makes use of the Kelvin point-source solution used by Horne1 in his treatment of the pressure build-up in wells. Use is also made of the super-position theorem. MATHEMATICAL DERIVATION The well is assumed to have been in production at a constant rate for a relatively long time (several days or more). It is then shut-in for a short time, after which it is again placed in production at the previous constant rate. It continues to produce at this rate until shut-in for a pressure build-up test. The Kelvin point-source solution is expressed by Eq. 1. The sequence of production rates is listed in Table I. TABLE 1 — SEQUENCE OF PRODUCTION RATES Time interval Production Rate At time AT the pressure will be the resultant of the sequence of production rates shown in Table 1. An expression for the super-posed effects may be derived from this table. (Effect of + q for T + ? T) + (Effect of - q for T + ?T-t,) + (Effect of + q for T + ? T - t,) + (Effect of - q for ? T) = Resultant........(2) The desired result is obtained by substituting the effect defined by Eq. 1 for the terms in Eq. 2. The result is Eq. 3. When the logarithmic approximation is made for the exponential integrals and the terms collected the result is Eq. 4. Let T — t2 = T = flowing time following short term shut-in t2 — t1 = 6 = duration of short term shut-in Then T - t1 = r + 6 Substitution of these quantities in Eq. 4 gives p(rw ?T) = P,, -When d<<t - AT then d—&apos;-, T— approaches 1. r + AT In this case the effect of the short term shut-in may be ignored. In order to get some idea of the numerical magnitudes involved take the worst possible case, when AT = 0, that is at shut-in. If the error is made less than some maximum acceptable value at this time, it will be less than the maximum for all subsequent times. d + r Let the difference between the value of y-— and 1 be less than 0.1. Then —6+r — 1 < 0.1 S < O.l T Therefore, if the flowing time following a short term shut-in is at least 10 times the duration of the short term shut-in, the error introduced by ignoring the short term shut-in will not exceed 10 per cent of T+AT A T

Jan 1, 1957

By B. Haimson, C. Fairhurst

A criterion is Proposed for the initiation of vertical hydraulic fracturing taking into consideration the three stress fields around the wellbore. These fields arise from (1) nonhydrostatic regional stresses in earth, (2) the difference between the fluid pressure in the wellbore and the formation fluid pressure and (3) the radial fluid flow through porous rock from the wellbore into the formation due to this pressure difference. The wellbore fluid pressure required to initiate a fracture (assuming elastic rock and a smooth wellbore wall) is a function of the porous elastic constants of the rock, the two unequal horizontal principal regional stresses, the tensile strength of the rock and the formation fluid pressure. A constant injection rate will extend the fracture to a point where equilibrium is reached and then, to keep the fracture open, the pressure required is a function of the porous elastic constants of the rock, the component of the regional stress normal to the plane of the fracture, the formation fluid pressure and the dimensions of the crack. The same expression may also be used to estimate the vertical fracture width, provided all other variables are known. The derived equations for the initiation and extension pressures in vertical fracturing may be employed to solve for the two horizontal, regional, principal stresses in the rock. INTRODUCTION Well stimulation by hydraulic fracturing is a common practice today in the petroleum industry. However, this stimulation process is not a guaranteed success; hence, the deep interest shown by the petroleum companies in better understanding the mechanism that brings about rock fracturing, fracture extension and productivity increase. Geologists and mining people became interested in hydraulic fracturing from a different point of view: the method may possibly be employed to determine the magnitude and direction of the principal stresses of great depth. Numerous articles in past years have dealt with the theory of hydraulic fracturing, but they all seem to underestimate the effect of stresses around the wellbore due to penetration of some of the injected fluid into the porous formation.1-4 Excellent papers on stresses in porous materials due to fluid flow have been published5,6 but no real attempt has been made to show the effect of these stresses in the form of a more complete criterion for vertical hydraulic fracturing initiation and extension. This paper is such an attempt. ASSUMPTIONS It is assumed that rock in the oil-bearing formation is elastic, porous, isotropic and homogeneous. The formation is under a nonhydrostatic state of regional stress with one of the principal regional stresses acting parallel to the vertical axis of the wellbore. This assumption is justified in areas where rock formations do not dip at steep angles and where the surface of the earth is relatively flat. This vertical principal regional stress equals the pressure of the overlying rock, i.e., S33 = —pD, where S33 is the total vertical principal stress (positive for tension), p is average density of the overlying material and D is the depth of the point where S33 is calculated. The wellbore wall in the formation is considered to be smooth and circular in cross-section. The fluid flow through the porous elastic rock obeys Darcy&apos;s law. The whole medium is looked upon as an infinitely long cylinder with its axis along the axis of the wellbore. The radius of the cylinder is also very large. Over the range of depth at which the oil-bearing formation occurs, it will be assumed that any horizontal cross-section of the cylinder is subjected to the same stress distribution, and likewise that it will deform in the same manner. STATE OF STRESS PRIOR TO DRILLING OF WELLBORE At a depth D from the earth surface within the oil-bearing strata, it is assumed that the following are 1References given at end of paper.

By Louis A. Rosales, Kanji Ono

The anomalous behavior of hydrogen in iron is examined and a new model is proposed It is postulated that hydrogen single and di-interstitials are in equilibrium with each other. When the binding enthalpy of the di-interstitial, B, is assumed to be about 10 kcal per mole, the predicted temperature dependence of the solubility agrees reasonably well with the reported data. Effective diffusiuity is calculated assuming the di-interstitials to be immobile. Since the number of mobile single interstitials is greatly reduced at lower temperatures, the effective diffusivity is decreased. A good agreement in the temperature dependence of diffusivity between theory and experiment is provided with B - 10 kcal per mole. The predicted diffusivity is also a function of total concentration of hydrogen and this appears to be Partly responsible for the large scatter in the reported data. HYDROGEN is known to form an interstitial solid solution with iron.&apos;&apos;&apos; At elevated temperatures, it has been well-established that hydrogen enters iron in the dissociated form and occupies the interstices, and that diffusion of hydrogen in iron is via the interstitial mechanism. The concentration, c, permeability, P, and diffusivity, D, of hydrogen in iron have been separately determined and are consistent with each other under the steady-state condition for temperatures above approximately 400°C; that is, the relationship P = c-D holds. At lower temperatures, however, anomalous behavior of hydrogen in iron has been observed by many investigators. The results of experimental studies4-&apos;&apos; of the diffusion of hydrogen in iron are summarized in Fig. 1, in which the values of D are plotted against the reciprocal temperature. Although there is an increasing divergence in observed values of D with decreasing temperature, the values of D at temperatures below approximately 200°C are definitely small as compared with the extrapolation from high-temperature data. The lines shown bracket the region containing the most probable data. The observed values of c6," are plotted against the reciprocal temperature in Fig. 2, where the dashed lines represent Geller and sun&apos;s&apos;&apos; most probable isobars at 1 and 100 atm. At high external hydrogen pressure and low temperatures, the values of c observed are higher than those expected from high temperature data. It is generally accepted that certain trapping mechanism is responsible for the anomalies, but the nature of the trapping mechanism has not been clarified. Various types of hydrogen traps were suggested in the literature, such as "rifts" in the lattice dislocations, voids, microcracks, and so forth. However, no quantitative account of the observed behavior has been given in terms of these trapping mechanisms. Formation of methane in hydrogen-charged steel was found by ~odgurski." This appears to explain the existence of the so-called "residual" hydrogen,4 but not the observed anomalies of hydrogen in iron. McNabb and Foster&apos; proposed a mathematical analysis in terms of a trapping mechanism. However, the physical significance of the traps was not clarified. The necessity of evaluating five parameters in their theory made comparison with experiment difficult. In the present paper, a model is proposed in which the anomalous behavior of hydrogen is attributed to the formation of hydrogen di-interstitials at lower temperatures. In this model, it is postulated that single and di-interstitials are in equilibrium with each other. (The hydrogen di-interstitial is to be distinguished from a hydrogen molecule which may exist in voids.) It can then be shown that most of the anomalies can be explained by assuming that the di-interstitials are virtually immobile and that the binding enthalpy of the di-interstitial, B, is approximately 10 kcal per mole. Predictions of the theory agree with the available experimental data reasonably well. THEORY The basic postulate of the present analysis is that single and di-interstitials are in equilibrium with each

Jan 1, 1969

By E. D. McMurry

INTRODUCTION Before discussing the use of the free piston; it may be well to explain briefly the type of gas lift valves with which the loo] is designed to function. The gas lift valve ii nothing more or less than a small back pressure regulator. It; construction consists of an outer hous ing covering a seamless metallic bellows in such a way as in form a sealed pressure vessel or pressure dome. To the bellow* in attached a stem and to this stem is attached a .rating member. When the valve is run into the hole, a pre-delermined pressure charge is placed in the pressure dome. As shown in Fig. 1. the valve is furnished in four different major assem-blies. to accomplish good results in different types of wells. The valve is mounted on a pup joint (Fig. 2a) : the eoin-plele assembly is entered into the tubing siring at the desired location. Each valve is furnished with a revere flow check valve (Fig. 2-b) to eliminate any back wash through the mechanism. Valve action is as follows: At any time the casing pressure is allowed to exceed the pressure forces within the pressure dome of the valve, the valve simply opens and passes gas from the rasing annulus into the tubing. When the casing pressure forces arc allowed to bleed to a force less than the pressure force within the dome, of the valve. the valve closes. With this type of construction. it is possible to open and close the valve or valves at will from the surface, the medium of the gas pressure in the rasing annnlus. The cycle of operation is governed by a tine cycle motor valve, commonly referred to as a gas lift interutitter or gas life surfacr controller (Fig. 3). In actual practice. after the well is unloaded only one valve is in operation . . . all other valve.: remain in the closed position. On pre-determined time cycles, it is possible to open and close this operating valve at will, and to pass the desired amount of gas through this valve wheneler needed. This method of gas lift is widelv used and has been found In be very efficient. However, it has long been known that even greater efficiency would be obtained if it were possible to place Piston between the injected gas and the fluid head. The free piston which will be discussed here was designed to accomplish this end and has been found to do so with very satisfactory results. DESIGN AND OPERATION OF THE FREE PISTON Basically. the free piston consist- of a telescoping assembly in which are incorporated the various control and operating parts (Fig. A set of sleeve valves which are opened and closed by movement of the telescoping section is contained in the. tool. A cross section of the free piston is shown in Fig. 4-11, &apos;rile main body cotain- a seamless monel bellows of the same type as that generally used in pressure charged gas lift valves (Fig. 4-c), and a valve core to permit placing a pressure charge in the bellows before the free piston is placed in service. The bellows is the main control unit of the free piston and performs two function: 1. It governs the opening and closing of the sleeve valves. a.: dictated by the producing activities of the well in which the piston is operating. 2. In its normal position, it holds a rubber packer in an elongated position to permit passage of fluid between the packer and the tubing walls. The lower section of the free piston contains a hydraulic shock absorber which cushions the return of the tool to its

Jan 1, 1953

By J. R. Lewis, J. H. Hamilton, V. D. Hanson, K. W. Heinlein, J. M.194-000-000-006 Vonfeld, J. P. Denny

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion NO. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and — -in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3h-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 34 in., and 1&apos;/4x in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into YSXO which is cleaned in the fine coal plant and 3x for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1953

By J. H. Hamilton, J. R. Lewis, K. W. Heinlein, V. D. Hanson, J. M. 194-000-000-006 Vonfeld, J. P. Denny

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion NO. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and — -in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3h-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 34 in., and 1&apos;/4x in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into YSXO which is cleaned in the fine coal plant and 3x for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1953

By W. W. West

MACHINERY used for size reduction of materials may be divided into three classes-depending upon the manner in which the crushing force is applied. In the first class could be listed the primary crushing units, such as the jaw crusher and gyratory crusher-which crush to coarse sizes by applying pressure intermittently between converging surfaces. In the second class may be listed crushing units, such as the hammer mill, roller mill, and ball mill, which crush principally by grinding or attrition. In the third class is the impact mill which, as the name infers, crushes by impact. The second and third classes generally include machinery used as secondary units for intermediate or fine crushing and grinding. Theory of Crushing by Impact Throw a stone into the air and give it a severe blow while it is momentarily suspended, and it shatters into several pieces. The impact crusher uses this "batting" principle, but carries it one step further. As the material to be reduced enters the crushing chamber, it receives a severe free-air impact blow by hard-hitting beaters suspended from rotor discs. The shattered particles formed are then driven at high velocity against impact blocks, where a further reduction in particle size takes place. The rebounding particles from the impact blocks are again caught by the beaters, and the process repeated until the material is swept out of the open bottom of the crusher. Impact Crushing Versus Attrition Crushing Fig. 1 shows a Hammermill type of attrition crusher. This particular mill has central feed and is reversible. Because of the central feed, an impact section can be designed into it at the feed entrance for preliminary crushing, followed by the cage bar attrition area for final sizing. The material is caught and held in a crushing chamber enclosed by grate bars or perforated plate. It remains in this chamber until it is reduced to particle size small enough to pass through the openings between the bars or in the plate. In this manner the top particle size discharged from the mill is closely controlled. As the swinging hammers pass through the bed of material, the material is broken down by a scrubbing action between hammers and material, material and grate bars, and material against material. This attrition produces a rounded particle shape, accompanied by slivers and a comparatively high percentage of fines. The crusher is especially useful in industries where the material treated is not too abrasive, such as limestone, cement or coal plants. It can be readily understood that from the maintenance standpoint, it is not suitable for hard abrasive materials such as refractories and aluminum oxides. One common type of impactor is shown in Fig. 2. The bottom of this type crusher is entirely open and the passage of the material through the crushing chamber is almost instantaneous. Liberal clearances are maintained between the impact hammers and the anvil type breaker blocks. This eliminates practically all attrition. It is the elimination of attrition that makes the impactor particularly adaptable for reducing hard, abrasive materials. The impactor is also well suited for the reduction of wet or frozen materials, which clog an ordinary Hammermill. Particles broken by impact, when examined under the microscope, show a distinct tendency toward rough, cubical shape, with a minimum of fines. Another important factor in impact crushing is that crushing is accomplished along natural cleavage lines, thus freeing foreign ores or materials.

Jan 1, 1952

By L. H. Sjodahl, J. B. Conway

of each particle on grain boundary movement is proportional to the area it occupies on the grain boundary and this increases with the square of the particle radius. In contrast to the impression gained by Tiwari the effectiveness of a particle, as a grain boundary obstacle, increases rapidly as the particle size increases. The counterbalancing factor in a dispersion of constant volume (e.g., those considered by Zener) is of course the rapidly diminishing number of particles, a situation which does not exist in thoria dispersed materials where the volume fraction of thoria remains constant. Although it has been assumed that the Zener approach is valid for the sake of the above argument, it is open to the following objections: 1) It assumes that the effect of many small particles along a grain boundary is cumulative and does not take into account the possibility of each small particle being overcome separately. In this case the boundary will not be able to move forward until the largest particle on the boundary has been overcome. This means that the rate controlling factor will be the rate at which a few relatively large particles are overcome with the remaining small particles playing a negligible role. 2) When the dispersion consists of very small particles the interparticle spacing is usually small compared to the grain size and most particles are located away from grain boundaries where they can play no part. The most efficient dispersion, from a purely geometrical viewpoint, is one where the particle spacing is equal to the grain size so that all the particles are on grain boundaries. Tiwari&apos;s second point that voids cannot prevent re-crystallization seems to be unfounded as several workes20-23 have shown that voids alone can retard grain boundary movement and recrystallization. To further convince Tiwari, Fig. 17 shows a void in deformed TD Nichrome holding back a recrystallization interface. Tiwari&apos;s explanation for the effect of voids on recrystallization, i.e., they reduce local strain energy and hence lower the driving force for recrystallization may be valid in itself but is irrelevant to the present case where even complete recrystallization after ONE of the more recently proposed expressions for describing stress-strain behavior is that due to Voce.1,2 This equation is usually written: s =s8 ?(s8?so)e??/?c [l] where s is the true stress, ? is true strain, cC is a constant called the characteristic strain, s8 is the asymptotic stress approached at high strain values, and so is the threshold stress at which the plastic deformation begins. While this expression has not been universally accepted as yet it does contain certain features which are consistent with experimental stress-strain behavior. In a more recent publication Hsu, Davies, and Royles3 employed the concept of generalized strain proposed by seth4 and identified a method for linearizing the stress-strain plot. In this approach the generalized strain was given by: where 7 is the generalized strain, 1 is the instantaneous length, 1, the datum length, and n is a constant called the "coefficient of strain measure". A value of n was said to exist which would yield the linearity: s —s+ S[.-(tf] [3] In recent studies of both these approaches, it has been found that while the Voce equation and the gen-

Jan 1, 1970

By Georges Caraghiaur

INTRODUCTION A small, independent mining enterprise (SIME) is an independently owned and operated business concern engaged primarily in mineral development, from the exploration to the production stage. By definition, a SIME is not dominant in its field and has no long-term financial backing. It is not a major producer of minerals, nor is it a petroleum company, a government organization or a subsidiary of such organizations. A SIME may or may not be listed on stock exchanges. The SIME is also called small miner, junior mining company, speculative mining company, and penny-stock mining company. The survival and growth of a SIME is most often controlled by an independently minded entrepreneur, who not only conceives and carries out new projects, but is materially affected by the success or failure of these projects. Usually, a SIME does not have a large or constant cash flow. Such a lack of long-term financial backing forces the SIME to meet the challenge of finding capital for its activities on a step-by-step basis. Its survival strategy is thus based on finding funds at the right time, of the right type and in the right quantities. The purpose of this paper is to show that the type of venture in which a SIME is involved affects the difficulty of raising new capital and the choice of possible sources of funds. The Small, Independent Mining Enterprise (SIME) The SIME is not a figment of the imagination of financial speculators. In Canada, there are more than 1,800 small mineral enterprises which spend on the average less than $1,000,000 per year on activity which is directed at bringing forth new mineral supply. In the United States, 64% of the non-fuel mining operations are conducted by nine or fewer men, and the coal industry is characterized by several thousand smaller independent producers, still predominantly family owned or controlled. In Chile, 80% of the mines can be classified as small mines. In the Philippines, approximately 60% of metal mining operations are classified as small mines. The SIME is actively engaged in mining all over the world, and very much so in the Pacific Rim. The Mineral Supply Process The whole process of bringing forth new mineral supply involves a dynamic sequence of activities, ranging from the initial exploration for a mineral occurrence to the production of a marketable product. A typical sequence is outlined in Figure 1. At the grass-roots exploration stage, a series of geological, geophysical and geochemical tests is aimed at locating a mineral occurrence. Especially in the early phases of such activities, the uncertainty (risk) surrounding a venture is high, since only general geological information exists on which to base a judgement of the economic viability of the venture. Such risk is attractive to few investors. As the information-gathering process continues, uncertainty diminishes. Once exploration efforts result in the discovery of a mineral occurrence, i.e. when indications of mineralization and mineable widths are obtained by drilling, there is still uncertainty concerning the extent of mineralization and the probability of having located an economic mineral deposit. The expected return on investment at such a stage may well be negative but the slight possibility of a very high positive return will make a number of financial sources available to an exploration company. At the development stage, much of the uncertainty

Jan 1, 1985

By W. M. McCardell, W. O. Winsauer

The abnormal conductivity found in shaly reservoir rocks containing an electrolyte is shown to be a consequence of the electrical double layer in the solution adjacent to charged clay surfaces. This increased conductivity results from a higher concentration of ions in the double layer than in the solution in equilibrium with the double layer. It is shown that the magnitude of the increased conductivity of a shaly reservoir material is influenced by the concentration and type of ions in the equilibrium solution as well as by the colloidal nature of the rock. INTRODUCTION An important factor in the quantitative interpretation of the electric log is the resistivity factor of reservoir rock. The resistivity factor is defined as the resistivity of the rock when completely saturated with an electrolyte divided by the resistivity of the electrolyte itself.&apos; In the normal case, the resistivity factor for a particular rock sample is independent of the resistivity of the electrolyte and reflects the pore geometry of the rock.&apos; In 1949. Patnode and Wyllie&apos; showed that in some cases the resistivity factor is not constant, but instead varies with the resistivity of the electrolytic solution. It was shown that clean sands behave normally—i.e., that the resistivity factor does not vary with resistivity of the electrolyte — but that sands con- taining shale and clay may exhibit abnormally low resistivity factors when the solution used to saturate them is of fairly high resistivity. At low resistivities of the electrolyte, the resistivity factor for a shaly sand appeared to approach a normal behavior. It is evident that the resistivity factor of a shaly sand is dependent upon factors other than pore geometry when the solution used to saturate it is of high resistivity. Patnode and Wyllie, and later de Witte,&apos; assumed that the conductivity of a shaly sand saturated with an electrolyte could be represented as the sum of two quantities. One of these was assumed, in effect, to be the conductivity which would be expected if the sample were a clean sand. and the other was assumed to be a conductivity inherent in the sample itself. The latter was supposed to be constant for a given rock sample regardless of the solution used to saturate it. The source of this added conductivity was ascribed to "conductive solids." In actual practice: the effect of the abnormal behavior of shaly sands is of minor importance except when the resistivity of a formation such as a sand is high. Thus the effect is more important when the sand is saturated with a dilute brine than when it contains a more concentrated electrolyte. The effect is also more important when the sample contains oil or gas than when it contains only saline water. Recently it has been suggested that the porosity of a stratum may be estimated from the electric log by use of the resistivity of the invaded zone." This is the zone flushed with the filtrate from mud used in drilling the well. which is ordinarily very dilute brine. When the method is applied to shaly sands, the effect of abnormal conductivity becomes of importance. This report describes an investigation of the mechanism responsible for the abnormal behavior of the resistivity of shaly rock saturated with conducting solutions. The investigation

Jan 1, 1953

By J. M. Vonfeld, V. D. Hanson, W. K. Heinlein

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion No. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and - 3/8-in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3/8-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 1/4 in., and 11/4x3/8 in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into 3/8x0 which is cleaned in the fine coal plant and 3x3/8 for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1952

By Thomas S. West

A gravel pack completion is described which, in addition to sand screening, permits the exclusion of water and gas from the oil production. A gravel pack composed of alternate layers of different particle sizes is employed which has horizontal but not overall vertical permeability to drilling mud. A packer is run on the tubing and set in the screen member, around which the gravel layers are placed. The hydrostatic pressure of a mud column prevents the entry of fluid above (or below) the level of the packer. The producing level may be changed by raising or lowering the tubing. Vertical coning is controlled by replacing the mud column with oil and continually injecting oil into the sand section above (or below) the packer. Three well histories with this completion are described in detail. Statistics on the first 100 completions of this general type are given. INTRODUCTION Gravel pack type completions have heretofore been almost exclusively for the purpose of preventing sand production. This paper describes and presents well data on a gravel pack procedure which is primarily a selective completion method. Although complete sand screening is achieved, the exclusion of gas and water is the primary objective. The application of this procedure is therefore not limited to unconsolidated sands. Application to hard sand or limestone reservoirs is equally feasible. GRAVEL PACK PROCEDURE Figs. 1 and 2 illustrate the gravel pack procedure employed. The gravel is placed by means of the frangible disc type dump bailer shown by Fig. 1.&apos; This bailer has its lower end closed by a plate glass or steel rimmed cement disc. This disc breaks when the bailer passes over the liner allowing the gravel to be deposited around the liner. The liner employed is usually 2 3/8 in. to 2 7/8 in. OD and 1.995 in. or 2.07 in. ID (ID of 2-in. tubing or 2-in. line pipe). Two tvDes,. of slotted sections have been employed. One type employs horizontal hacksaw cut or milled slots of about 1/32 in. width opposite the producing section. The other type employs the same type,. of horizontal slot. but slots are placed in groups of six at intervals of about 8 in. A cage composed of half round vertical bars, spaced apart, is then welded to the pipe over the slotted sections. This liner has the bottom end closed, usually by bull plugging. The upper end is closed, while placing the gravel, by a cap which is held in place by means of a shear pin. This cap has a mandrel attached to it which extends downward inside of the liner to within a few inches of its lower end. The top of the liner cap is formed to fit a releasing type wire line core barrel overshot. The liner is lowered to bottom on the wire line by means of this overshot and released. The length of the liner is usually suffi- cient to extend to about 20 ft above the highest level to which gravel is to be placed, the upper end of the liner always being in the casing. The dimensions of the liner and bailer are such with respect to the casing size that the bailer will always pass over the liner. After releasing the liner the bailer is run to bottom. Passage of the empty

Jan 1, 1951

By Thomas S. West

A gravel pack completion is described which, in addition to sand screening, permits the exclusion of water and gas from the oil production. A gravel pack composed of alternate layers of different particle sizes is employed which has horizontal but not overall vertical permeability to drilling mud. A packer is run on the tubing and set in the screen member, around which the gravel layers are placed. The hydrostatic pressure of a mud column prevents the entry of fluid above (or below) the level of the packer. The producing level may be changed by raising or lowering the tubing. Vertical coning is controlled by replacing the mud column with oil and continually injecting oil into the sand section above (or below) the packer. Three well histories with this completion are described in detail. Statistics on the first 100 completions of this general type are given. INTRODUCTION Gravel pack type completions have heretofore been almost exclusively for the purpose of preventing sand production. This paper describes and presents well data on a gravel pack procedure which is primarily a selective completion method. Although complete sand screening is achieved, the exclusion of gas and water is the primary objective. The application of this procedure is therefore not limited to unconsolidated sands. Application to hard sand or limestone reservoirs is equally feasible. GRAVEL PACK PROCEDURE Figs. 1 and 2 illustrate the gravel pack procedure employed. The gravel is placed by means of the frangible disc type dump bailer shown by Fig. 1.&apos; This bailer has its lower end closed by a plate glass or steel rimmed cement disc. This disc breaks when the bailer passes over the liner allowing the gravel to be deposited around the liner. The liner employed is usually 2 3/8 in. to 2 7/8 in. OD and 1.995 in. or 2.07 in. ID (ID of 2-in. tubing or 2-in. line pipe). Two tvDes,. of slotted sections have been employed. One type employs horizontal hacksaw cut or milled slots of about 1/32 in. width opposite the producing section. The other type employs the same type,. of horizontal slot. but slots are placed in groups of six at intervals of about 8 in. A cage composed of half round vertical bars, spaced apart, is then welded to the pipe over the slotted sections. This liner has the bottom end closed, usually by bull plugging. The upper end is closed, while placing the gravel, by a cap which is held in place by means of a shear pin. This cap has a mandrel attached to it which extends downward inside of the liner to within a few inches of its lower end. The top of the liner cap is formed to fit a releasing type wire line core barrel overshot. The liner is lowered to bottom on the wire line by means of this overshot and released. The length of the liner is usually suffi- cient to extend to about 20 ft above the highest level to which gravel is to be placed, the upper end of the liner always being in the casing. The dimensions of the liner and bailer are such with respect to the casing size that the bailer will always pass over the liner. After releasing the liner the bailer is run to bottom. Passage of the empty

Jan 1, 1951

By M. H. Gaskell, D. M. Kehn, G. T. Pyndus

A correlation of the bubble point pressure for black oil systems is developed using the standard physical-chemical equations of soiutions. The correlation is based on 158 experimentally measured bubble point pressures of 137 independent systems and is expressed in terins of the usually measured field pnrameters— flash separation gm-oil ratio, rank oil gravity, total gas gravity, and reservoir temperature. The data were obtained on systems produced in Canada, Western and Mid-Continent01 United States, and South America. The average error (algebraic) in the representation is 3.8 per cent, and the maximum error encountered is 14.7 per cent. INTRODUCTION In the absence of experimentally measured properties of reservoir fluids, it is often necessary for the field engineer to make estimates regarding the fluid properties based on the usually measured producing parameters. To aid in these estimations, various correlations have appeared in the literature in recent years. Among the pertinent properties of interest is the bubble point pressure. A correlation for this parameter has been reported by Standing&apos;. However, this correlation was based essentially on California produced crudes and this limitation was pointed out with its presentation. The correlation presented in this paper utilized data on crude oil systems from Canada, Wes- tern and Mid-Continental United States, and South America. CORRELATION DEVELOPMENT The basic assumption used in this development is the same as employed by Standing&apos;, There is a wide variety of ways to combine these parameters; however, in this instance the combination was made on the basis of Henry&apos;s law2. Accordingly, the relationship proposed is Although Eq. 2 defines an individual system, it is of limited value since H&apos; is a function of gas-phase composition and the system temperature. It was observed that for the systems where the bubble point was measured at several temperatures that the ratio of the bubble point pressures and the ratio of the corresponding absolute temperatures (R) were practically identical. Thus, for correlation purposes the bubble point pressure may be taken as a direct function of the absolute temperature. This relation is valid only for those systems that are not near the critical point. Accordingly, this correlation will be inadequate for systems in the region of the critical point. The solubility of the various hydrocarbons found in the gas phase increases with the molecular weight. Thus, the saturation pressure should be inversely related to the gas gravity. Applying these principles to Eq. 2 and rearranging terms gives: The variables 02 the left side of Eq. 3 were designated as the "hubble point pressure factor". The number of mols of tank oil per barrel is a function of the "mclecular weight" of the tank oil. Although the tank oil is a complex mixture, it was assumed for the purposes of this correlation that a unique molecular weight could be assigned to a given oil. This was designated as the "effective molecular weight", and was related to the oil gravity, This empirical relationship was developed simultaneously with the correlation by assuming values of M,. and working to obtain a smooth curve for both the correlation and the effective molecular weight. The relationship between the oil gravity and the effective molecular weight used in this correlation is shown in Fig. 1. The effective molecular weight is somewhat higher than the molecular weight of the C fraction. The difference between these values is largest for the low-gravity systems. It is noted that this effective molecular

By H. H. Waller, D. L. Monroe, D. P. R. Smyth

ORIGINALLY slusher drift development was conventional, advancing the drift full 10x13-ft size at 6 ft per round. This proved dangerous and costly because the weak fractured rock of the orebody cannot be left raw for any length of time without its reacting with air to cause spalling. With this method of drift development ground conditions were worsened by many blasts and the length of time it was necessary for the drift to remain open. The problem was practically solved by driving 5x7-ft pilot drifts, then enlarging to 10x13-ft raw size by slashing with 50 to 60-ft diamond drill holes. Calculated explosive charges are used to avoid wall and roof shattering. Still more recently jacklegs and sectional steel replaced the diamond drills, resulting in the same work for less than half the cost, see Figs. 1 and 2. Crosscuts at right angles to the main haulage drifts, which cross through the center of the ore-body, divide the ore into 200-ft blocks. Eight slusher drifts on 50-ft centers draw from one block. The slusher floor is 11 ft above the rail, inclined at 1 pct grade, and the top is 18 ft below the undercut. The finished slusher drift is 102.5 or 120 ft long, depending on whether it is north or south of the crosscut. Drawpoints are staggered on each side at 17.5-ft centers. The completed drift is concreted and has an inclined ramp and a vent at the back. When crosscuts and gangways have been excavated and concreted and slusher stations installed, slusher drift development is carried out in 10 clear-cut stages: 1—Pilot drift is driven full length, 5x7 ft. 2—Station is excavated at half-way point in pilot drift. 3—Pilot drift is enlarged to slusher drift size by drilling. 4—Lower part of slash is blasted and mucked for floor concrete. 5—Floor is poured. 6—Inside upper half of slusher drift is blasted, mucked, and concreted. 7—Outside upper half of slusher drift is blasted, mucked, and concreted. 8—Drawpoints are excavated. 9—Drawpoints are concreted. 10—Concrete ramp and sheave protection recess are constructed at back end of slusher drift. A 5x7-ft pilot drift is driven full length on the center line of the proposed slusher drift using a jackleg and detachable bits on 7/8-in. hexagon steel. A 10-hp air-driven slusher and 36-in. scraper pull muck directly into cars in the main haulage crosscut, see Fig. 2. Operations are on a three-shift basis. One man mucks the previous round, drills, and blasts his round in a shift. Average advance is 5 ft per shift. When the pilot drift is completed, 7 ft of drift in the center section is enlarged to full raw slusher drift size of 10x13 ft, for the set-up to drill long slash holes. The pilot drift is slashed to full raw size from this station with jacklegs and tungsten carbide chisel bits on 7/8-in. sectional steel. Crews line the holes without trouble, working from controls established by the engineers, and the pattern of holes for the slash is arranged to give the least shattering effect

Jan 1, 1955

By J. D. Meakin

J. D. Meakin (The Franklin Institute Laboratories)— In a recent paper Garofalo12 has shown that a number of experimental creep results can be represented by the empirical relation In this expression ? is the steady creep rate, s is the applied stress, and A is a constant at a given temperature; a and n are adjustable parameters. Garofalo has shown that Expression [6] approximates to a power function at low stresses and a simple exponential function at high stresses. Among the data used by Garofalo are the results we published for creep of copper." For these results we found that the best correlation was an exponential law with different parameters in the low-and high-stress region, respectively. That is, i = A exp qs where A and q have different values above and below a critical stress a,. In Ref. 13 are listed a variety of observations which support the proposal that different mechanisms are operative at stresses above and below a,, and it is not felt appropriate to repeat those observations here. The purpose of this note is to demonstrate that, although the representation used by Garofalo does fit the experimental data very well, this fit cannot be used to conclude that only one mechanism is rate-controlling. The argument is best made graphically and is illustrated in Fig. 9. The continuous curves 2 and 3 represent pairs of simple exponential functions which intersect at i = 10-5. The curve 2 closely resembles the copper results in that a change of slope of 2:1 occurs at i = 10-5; the two functions shown are in fact: e = 10-7 exp (1.150) and ? = 10-6&apos; exp (0.576s) For purposes of comparison, curve 3 represents two exponential functions with a change of exponent of 3:1 and curve 1 has the exponent 1.150 throughout. It is quite clear that a satisfactory straight-line fit is obtained on a Garofalo plot for all three curves. The curves represent processes which have either a single stress exponent or two ex- ponents in a ratio as high as 3:l. It must, therefore, be concluded that the ability to obtain a linear dependence between log i and log (sinh cyo) cannot be taken as evidence for a single activated process. In fact, to some extent the way the experimental points quoted by Garofalo lie about a best fit straight line on the log (sinh as) plot indicates that two processes are indeed occurring. This is seen in Figs. 7(c) and 8(c) of Ref. 12 where it is particularly evident that the experimental points deviate from the straight line in a manner similar to curves 2 and 3 of Fig. 9. It is, of course, recognized that the generally used exponential form is an approximation to the hyperbolic sine function which is not valid for exponents of less than 1.2, e.g., qs < 1.2. The only curve in Ref. 12 which is affected by this restriction is Fig. 6(b) where the deviation from a In ? vs s

Jan 1, 1964

By E. H. Crabtree, Neil S. Parker

THE custom milling plant of the Eagle-Picher Mining and Smelting Co. is at Sahuarita, Arizona, approximately 20 miles south of Tucson. It is connected by a 2-mile railroad spur to the main line of the Southern Pacific Railroad between Tucson and Nogales. The capacity of the plant is 500 tons per day. The mill is conventional except perhaps in the rather wide variety of ores treated. These vary from straight copper ores, from which only one concentrate is produced, and lead-zinc or copper-zinc ores from which two concentrates are produced, to complex gold-silver-copper-lead-zinc ores, which require the simultaneous production of three or four concentrates. Custom ores are received in lots varying in size from a few truckloads to several carloads. Similar ores, after crushing and sampling, may be commingled, or individual lots may be milled individually. The ores are first dumped either into one of two 200-ton railroad bins, or into one of the two 75-ton truck bins. From these bins the lot of ore is individually crushed in a 24 by 16-in. jaw crusher to 3-in. size, and then to 1/2-in. size in a 3-ft Symons shorthead crusher. After crushing to 1/2-in., the ore is conveyed to the automatic sampling plant. This consists of three Vezin samplers in series, each successively cutting out 10 pct, to pct and 5 pct of its feed. After No. 1 sampler, the material passes through a mixing barrel and is then crushed to 1/4-in. in a set of rolls before passing to No. 2 sampler. Between No. 2 and No. 3 samplers, the ore is again mixed in a mixing barrel. The final sample, consisting of one pound per ton of ore, is taken to the laboratory, where it is further crushed to 1/8-in. in a coffeemill, and then riffled down to the size required for grinding in a pulverizer. The capacity of the crushing and sampling plant is up to too tons per hour, depending upon the character of the ore. After sampling, the ore is stored in five ore-storage bins, consisting of four 200-ton bins and one 175-ton bin. Each of these bins is discharged by means of belt feeders and conveyor to the ball mills, so that ores may be composited for milling in any required ratio, or milled separately by themselves. Fine grinding is done in one 6 by 4-ft Allis-Chalmers ball mill and one 8-ft by 36-in. Hardinge mill. These two mills both discharge to one 54-in. Akins Simplex Highweir classifier. The classifier sands can be returned to either or both mills, so that flexibility in grinding for different ores is easily obtained. Flotation equipment consists of four 66-in. Fagergren machines for copper-lead roughing and six similar machines for zinc roughing. Lead-copper bulk concentrate is cleaned and recleaned in eight No. 18-S Denver machines, and zinc concentrate is cleaned in three 66-in. Fagergrens. Copper-lead differential separation

Jan 1, 1947

By Louis Moyd

In the preparation of illustrations to accompany reports of investigations concerning particle shapes of various natural and manufactured materials proposed for use as he aggretates in concrete structures, it was found that stereoscopic views of such materials presented the information much more effectively than the usual two-dimensional photographs. With the whole-hearted cooperation of the Photographic Section, a technique was developed in the petro-graphic laboratory of the Concrete Research Division, U.S. Corps of Engineers, at Clinton, Miss., for simple and rapid preparation of paired stereoscopic photographs and micrographs. The method requires only equipment which is usually part of the general stock of any laboratory, and it is adaptable for all purposes in which three-dimensional views would be found superior to ordinary photographs. Stereoscopic Photographs Where the subject or subjects to be photographed are small but not of microscopic size, an ordinary camera having a ground-glass focusing screen is set up vertically over the field. The material to be photographed is placed on a card, the surface of which is suitable as a background. The card is mounted so that it can be tilted about 5&apos; above and below the horizontal plane, on an axis that coincides with the north-south diameter of the viewing screen. A simple way of achieving this is to mount the card on an object having a convex lower surface. In our laboratory, a watch-glass has been used, with plasticene as the mounting medium. The convex surface should be mounted on a support having a round opening on top and a flat bottom. A roll of tape, a petri dish, a beaker, or any other piece of equipment having the proper shape to give stability to the setup will be satisfactory. The arrangement of the required equipment is shown in Fig 1. For the best results, the subject to be photographed should be mounted in the plane of the axis of rotation. To achieve this effect, the depth of the mass of plasticene between the background and the watch-glass is varied so that when the watch-glass is tilted back and forth on the axis, the subject to be photographed does not appear to move back and forth across the viewing screen. A permanent piece of equipment, embodying an adjustable vertical screw, with lock-nuts, could be made for this purpose. The subject may be illuminated in any manner desired. It was found that the presence of shadows intensifies the effect of depth. In taking the photographs, the concave surface is mounted so that one side of the background card is about 5&apos; below the horizontal, while the axis remains in the horizontal plane. The subject is focused carefully on the ground glass and a picture is taken. The card is then tilted so that the opposite side is about 5&apos; below the horizontal, while the axis still remains in the horizontal plane. The subject is again put in focus and another picture is taken. After both pictures are developed and printed, they are placed side by side in the same orientation and examined with a stereoscopic viewer. If depressions appear where there ought

Jan 1, 1950

By E. H. Crabtree, Neil S. Parker

The custom milling plant of the Eagle-Picher Mining and Smelting Co. is at Sahuarita, Arizona, approximately 20 miles south of Tucson. It is connected by a 2-mile railroad spur to the main line of the Southern Pacific Railroad between Tucson and Nogales. The capacity of the plant is 500 tons per day. The mill is conventional except perhaps in the rather wide variety of ores treated. These vary from straight copper ores, from which only one concentrate is produced, and lead-zinc or copper-zinc ores from which two concentrates are produced, to complex gold-silver-copper-lead-zinc ores, which require the simultaneous production of three or lour concentrates. Custom ores are received in lots varying in size from a few truckloads to several carloads. Similar ores, after crushing and sampling, may be commingled, or individual lots may be milled individually. The ores are first dumped either into one of two 200-ton railroad bins, or into one of the two 75-ton truck bins. From these bins the lot of ore is individually crushed in a 24 by 16-in. jaw crusher to 3-in. size, and then to ½-in. size in a 3-ft Symons shorthead crusher. After crushing to ½-in., the ore is conveyed to the automatic sampling plant. This consists of three Vezin samplers in series, each successively cutting out 10 pct, 10 pct and 5 pct of its feed. After No. I sampler, the material passes through a mixing barrel and is then crushed to -in. in a set of rolls before passing to No. 2 sampler. Between No. 2 and No. 3 samplers, the ore is again mixed in a mixing barrel. The final sample, consisting of one pound per ton of ore, is taken to the laboratory, where it is further crushed to 1/8-in. in a coffeemill, and then riffled down to the size required for grinding in a pulverizer. The capacity of the crushing and sampling plant is up to 100 tons per hour, depending upon the character of the ore. After sampling, the ore is stored in five ore-storage bins, consisting of four 200-ton bins and one 175-ton bin. Each of these bins is discharged by means of belt feeders and conveyor to the ball mills, so that ores may be composited for milling in any required ratio, or milled separately by themselves. Fine grinding is done in one 6 by 4-ft Allis-Chalmers ball mill and one 8-ft by 36-in. Hardinge mill. These two mills both discharge to one 54-in. Akins Simplex High weir classifier. The classifier sands can be returned to either or both mills, so that flexibility in grinding for different ores is easily obtained. Flotation equipment consists of four 66-in. Fagergren machines for copper-lead roughing and six similar machines for zinc roughing. Lead-copper bulk concentrate is cleaned and recleaned in eight No. 18-S Denver machines, and zinc concentrate is cleaned in three 66-in. Fager-grens. Copper-lead differential separation

Jan 1, 1949