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By Ellis H. Gates

BENEFICIATION of the manganese oxide ores at Three Kids Mine near Henderson, Nev., has evolved over a period of years. Commercial application of the process is on a secure basis, and an effective working hypothesis has been formulated to help solve the problems encountered in flotation. An outstanding difference between the Manganese Inc. operation and other flotation processes is the necessity for high recovery of slime mineral values. Part of the manganese minerals in the Three Kids ore deposit occurs as primary slime or as minerals that are easily slimed by grinding. High recovery of this slime portion is essential to an economic overall manganese recovery. The aim of much of the research at Manganese Inc. was to recover these slimed and finely ground minerals by flotation. Early Agglomeration Processes: Tunbridge' in 1880, followed by Everson2 in 1886 and Cattermole" in 1904, discovered that under certain pulp conditions precious metals and sulfide minerals could be agglomerated if enough soap or oil were mixed into the pulp slurry. These agglomerated particles, Which under ideal conditions ranged from 1/16 to 1/4 in. diam, were recovered by screening, tabling, and hydraulic classification. Even flotation by air occluded in the agglomkrate particles by vigorous stirring was suggested. Cattermole nearly succeeded in getting his process into commercial operation, but it was finally abandoned, presumably because he could nat compete with the low reagent requirements of the new agitation froth process. Since Cattermole's time Christensen in 1921, Borcherdt4 in 1926, and Chapman and Littleford- in 1934 have attempted to apply the principles of agglomeration. In the light of current research at Manganese Inc. on conditioning of pulp at high energy input levels, it is interesting to note the variety of equipment such as tumbling mills, mixing drums, and shakers described by Borcherdt in his patent to produce the agglomerated particles. Ore Characteristics: Manganese in the Three Kids ore occurs as a 'complex mixture of manganese oxides commonly Known as wad. The manganese minerals have Been identified as psilorflelane, hollandite, coronadite, and pyrolusite with minor amdunts of rhodonite and rhodochrbsite. The ppincipal gangue miherals are quaktz, opal, kaolinite, montmoril-lonite, calcite, gypsum, fcelesitite, and barite. Most of the mangandse ore minerals are either porous and sponge-like in structure or soft clay-like masses composed of fine grained mineral particles approaching the size range of colloids. Consequently the ore minerals from the Three Kids mine have relatively large surface areas and high void content per unit weight. The specific area of various manganese ore samples measured at the California Research Corp. laboratories ranged from 32 to 58 sq m per g. The great surface area and high void content explains in part the unusually high reagent consumption required in the treatment of this ore. Current Milling Practice: At Henderson6 he ore is crushed to —1 in. and ground with rod mills in a closed circuit with spiral rake classifiers and cyclones. Sulfur dioxide solution is added to the pulp as it leaves the grinding circuit, immediately followed by addition of a reagent mixture of soap, oil, and wetting agent in the form of an emulsion. The reagents are initially well mixed into the pulp during their passage through centrifugal pumps to the conditioning equipment. All reagents are added prior to conditioning. The intense conditioning or mechanical stirring of the pulp, in which large amounts of power are consumed, is the principal difference between the flowsheet used at Manganese Inc. and a standard flotation circuit. As much as 38 kw-hr of energy per dry short ton of ore in a pulp containing 18 to 23 pct solids is used for conditioning. After conditioning, the pulp passes to a flotation circuit that consists of one roughing stage, a scav-

Jan 1, 1958

By D. Akers, R. Dospoy

Introduction The geological processes that form coal can also concentrate trace elements in the coal. For example, the average concentration of arsenic in bituminous coal (20 ppm) is ten times the average concentration found in all the other rocks that make up the earth's crust (2 ppm). Similarly, other elements, such as antimony, cadmium, mercury and selenium, are more concentrated in coal than in the earth's crust. When coal is burned, trace elements can be further concentrated. Although no new constraints on trace clement emissions were placed on the power generation industry under the 1990 Clean Air Act Amendments, the act does mandate a three-year study of air toxics. Any new regulations aimed at electric utilities as a result of this Federally-mandated study will almost assuredly be very costly-an estimated $1 billion per year. Many trace elements in coal are associated with mineral matter. For example, arsenic is commonly associated with pyrite, cadmium with sphalerite, chromium with clay minerals, mercury with pyrite and cinnabar, nickel with millerite, pyrite and other sulfides, and selenium with lead selenide, pyrite and other sulfides (Finkelman, 1980). There are also cases in which some of these elements are organically bound. Just as both organic and pyritic sulfur can be found in the same coal, the same trace element may be both organically bound and present as part of a mineral in the sane coal. Physical coal cleaning techniques are effective in removing mineral matter from coal and can potentially remove at least some of the trace elements associated with specific minerals, thereby reducing the release of these elements into the atmosphere. Conventional coal cleaning to remove trace elements As part of a project funded by the Electric Power Research Institute, CQ Inc., a wholly-owned EPRI subsidiary located in western Pennsylvania, has demonstrated that large reductions in the concentration of many trace elements are possible if conventional coal cleaning techniques are properly applied. Four examples are given in Tables 1 to 4. In each example, the results shown were generated by cleaning the coal at CQ Inc.'s commercial-scale cleaning test facility. Cleaning results for Upper Freeport Seam coal from Northern Appalachia are provided in Table 1. Data are presented in the table in two ways: as a weight-based concentration (parts per million) and as a concentration per heat unit (grams per billion Btu). Grams per billion Btu is analogous to pounds per million Btu, but avoids the use of numbers with many decimal places. The heat-based concentration provides a better measure of boiler impacts, because the increased heating value obtained through coal cleaning reduces the number of tons that must be burned to produce a given thermal output. Reducing the quantity of coal burned reduces the quantity of trace elements entering the boiler. This raw coal is relatively high in several trace elements of environmental concern, including arsenic, cadmium and chromium. Cleaning provided large reductions in the quantity of arsenic, barium, cadmium, chromium, fluorine, lead, mercury, nickel, silver and zinc. The results for tests with a Powder River Basin coal, Rosebud/McKay, are presented in Table 2. Large reductions in arsenic, barium, cadmium, fluorine, mercury, nickel, selenium and zinc were observed with cleaning. The concentration of chromium increased with cleaning, while lead concentration increased in one test and decreased in another. Table 3 presents test results for Croweburg Seam coal from Oklahoma. Large reductions in arsenic, barium, cadmium, chromium and zinc were obtained with cleaning. Smaller reductions were obtained with lead and nickel, while chromium, fluorine and mercury increased in at least one of the tests. Table 4 presents cleaning test data for Kentucky No.11 Seam coal. In this case, large reductions were obtained with all elements measured. In general, these data indicate that physical coal cleaning is effective in reducing the concentration of many trace elements, especially if they are present in the coal at relatively high concentrations. The degree of reduction achieved is coal-specific, relating in part to the degree of mineral association of the specific trace element and the degree of liberation of the trace element-bearing mineral. The extent of trace element removal also depends on the method of cleaning the coal. Figure 1 is a washability plot, by size fraction, of arsenic vs. ash content for Upper Freeport

Jan 1, 1994

By C. S. Smith, L. Guttman

It is shown, from a study of geometric probabilities, that the average number of intercepts per unit length of a random line drawn through a three-dimensional structure is exactly half the true ratio of surface to volume. Since the surfaces can be internal or external, the area of grain boundary or of the interface between any two constituents in a micro-structure can be measured. Other metric relations are tabulated that may be of use in studies of the microstructure of polycrystalline, cellular, or particulate matter generally. IN many fields of scientific investigation the structure of cellular aggregates or random arrays of discrete particles imbedded in some matrix is observed on a two-dimensional section and inferences are drawn therefrom as to the real structure in three dimensions. The biologist's microtome slice, the petrologist's thin section, and the metallurgist's plane polished and etched sections are common examples, although the problem is a general one. Scientists have commonly limited their thinking to the same dimensionality as their structures, and the few attempts that have been strictly three-dimensional in character have been laborious and noteworthy. From a metallurgical standpoint it is often of considerable importance to know, in addition to the volume fraction of two or more components in an alloy, the amount of two-dimensional interface between crystals. Such grain boundaries (which may separate either two identical crystals differing only in orientation or two crystals differing in structure, and possibly also in orientation) have a determining factor upon the mechanical behavior. It is at these boundaries that melting commences, that stress-induced corrosion occurs, and that various precipitates (harmful or otherwise) first appear. The boundary is doubtless of equal importance in nonmetallic crystalline aggregates such as rocks, ceramics, and concrete, and the biologist is deeply concerned with the area of cellular membranes. Many synthetic cellular foams involve similar structural problems. The very term structure usually implies the presence of interfaces and a complete understanding of structure involves nothing but an analysis of the geometrical, metrical, and topological relations between the various interfaces (zero, one and two-dimensional) that exist in a three-dimensional structure. Even systems lacking sharp physical interfaces often have interrelated gradients of composition or velocity (as cored crystals or turbulence cells) in which a neutral surface can be treated as a two-dimensional interface. In an earlier paper by one of the authors' the question of cell shape was considered in terms of simple topological principles without regard to physical dimensions. The determination of the actual size of grains in two dimensions is carried out in a routine fashion in innumerable metallurgical laboratories (see, for example, the ASTM standard methods of grain size determination2), though this is done merely to check the uniformity of a product and has no relation to the actual three-dimensional shape or size of the grains. Some authors have discussed the three-dimensional problem but only on the basis of assumptions as to idealized grain shapes.:'-' Quantitative measurements of microstruc-tures to obtain the volumetric relations of various phases have been carried out by petrographers for many years and are of increasing popularity among metallurgists." The present paper will show how, on the basis of no assumptions other than randomness of sectioning (usually realizable in experiment), it is possible to learn a great deal about the three-dimensional structure. The relations to be derived will generally be used on random arrays of cells or other particles, although they are equally applicable to ordered arrays and even to isolated objects of complex shape provided that suitable random sections can be made. Total Area of Interfaces in a Sample Consider a typical microstructure of a single-phase polyerystalline metal, such as that shown pres- in Fig. 1. The plane cross section shown contains a network of lines which subdivides the area into two-dimensional cells. The lines, of course, represent merely the intersection of the two-dimensional plane of sectioning with the two-dimensional interfaces between adjacent three-dimensional cells. The structure also contains points at which two or more lines intersect: in three dimensions the points are, of course, lines. In the more general case there may be in three dimensions isolated particles surrounded by a single interface without contact with others, and both two and one-dimensional features which do not necessarily connect with others. On the two-dimensional section these will appear as areas delineated by a closed line (as at a and b, Fig. 2), as isolated lines

Jan 1, 1954

By Fred Leighton

Miners have long known that rock noise, or the popping and cracking of the rock commonly heard during mining, can indicate instability of the mine structure. For many years, miners have "listened" to the rock talk and many times have interpreted changes in rock noise activity to be a warning of failure and have retreated from the potential failure area. Microseismics, or the study of rock noises, was begun in the early 1940s partly because of this historical fact. Microseismics is based on using geophysical equipment to detect and analyze rock noises on audible and subaudible levels. Thus, these systems are much more sensitive than the human ear and "hear" even more of the rock "talk" than do miners. Research has shown that micro-seismics can be used to precisely locate portions of a working area that are generating rock noise, and that rock noise release rate information from each area can be used to analyze its stability. On the basis of research results, microseismic systems are now commercially available and have been purchased and installed in many mines around the world. Introduction When a rock mass is subjected to changing stress conditions, such as those caused by mining, small-scale adjustments occur within the rock that release seismic energy. When this energy is in the audible range it is called rock noise. Those areas where stress changes occur are also the areas of the structure most likely to fail. Individual rock noises can be detected and analyzed to determine their precise location relative to the mine structure. Over a period of time plots of this data provide a pictorial representation of where stress changes are occurring. This is because rock noise activity tends to concentrate in those structure areas most actively adjusting to the changing stresses. Since these areas are the most likely to fail, they can be pinpointed and mapped relative to the structure. Rock noise rates, or the number of rock noises occurring per unit of time, also tend to vary dramatically before failure. Thus, the ability to locate the source of individual rock noises provides a means of determining where failure may occur, and rate counting within each area offers a means of assessing when failure may occur. This information, properly treated, can be used to avoid, control, or warn of impending failure. Successful applications of this technique have provided recent impetus to the effort of developing microseismics into a practical, reliable, and economically feasible tool. The phenomenon of naturally occurring rock noise was discovered in 1938 by Obert and Duvall (1939) who were measuring seismic velocities in mine pillars. Seismic energy other than what they were generating continually registered on their recording equipment. Further study by these researchers showed that the extraneous seismic energy being recorded was from rock noises that were being generated within the rock in highly stressed areas. Pursuing this interesting phenomena both in the laboratory and in the field, Obert and Duvall documented the dramatic change in rock noise rate before failure (Obert, 1941). They also established the fact that, in many instances, rock burst failures could be predicted by monitoring and listening to the rock noise activity in rock-burst-prone areas (Obert and Duvall, 1945). These early efforts clearly showed that microseismics had great potential as a tool for measuring or estimating mine structure stability. Extended testing, however, showed that sometimes rock bursts occurred with no apparent microseismic warning, and that at other times sharp increases in microseismic data were not accompanied by failure. Also, because precise location of individual rock noises was not possible at that time, one never knew where the failure was going to occur, only that failure near the observation point was likely. Thus, while the technique clearly offered promise, it was not considered practical then. In the mid-1960s, the Bureau of Mines undertook a new research effort to improve the microseismic technique. Major improvements were judged possible mainly because new and vastly improved electronic and system components had become available as an offshoot of space program instrument development. Thus, in 1967, development of a multichannel, broad band, microseismic system began (Blake and Leighton, 1970). Application of this system in rock-burst-prone mines showed it worked well and provided the incentive to develop methods whereby the source location of individual rock noises could be easily and directly calculated (Leighton and Blake, 1970; Leighton and Duvall, 1972; Redfern and Munson, 1982). The improved system and the ability to locate rock noises quickly showed that the microseismic technique offered new and increased poten-

Jan 8, 1983

By R. M. Genevray, M. B. Bever, E. J. Suoninen

THE martensitic transformation in the ß-phase of the Cu-Zn system has been the subject of several investigations. The transformation is known to be reversible and to be affected by stress. Its temperature range has been determined as a function of composition. In the investigation reported here, the effect of tensile stresses on the transformation was investigated quantitatively. Some information was also obtained on the thermoelastic behavior of the martensite formed in the first stages of the transformation. Most of the experiments were done with alloy E of an earlier investigation;' this alloy analyzed 60.49 pct Cu and 39.51 pet Zn by weight. The methods of shaping and heat treatment were also essentially the same as those previously used. The stress was applied to the specimen immersed in a cooling liquid. The transformation was followed by measuring the electrical resistance with a Kelvin bridge and the elongation with a cathetometer. Fig. 1 shows the M, temperature as a function of stress. Resistance and strain measurements gave essentially identical values. The results suggest a roughly linear relation between M. and in the range investigated, up to 12 kg s mm". At higher stresses, plastic deformation begins to interfere seriously with this relationship. The increase of M, with stress is consistent with published work on the effect of stress on the martensitic transformation. The slope of the curve, 4°C per kg mm ", is of the same order of magnitude as the corresponding value calculated for steel.' Fig. 1 also shows the difference, AM, between the temperature of 50 pct transformation on cooling, as measured by changes in length, and that of 50 pct reverse transformation on heating. This difference, which may be considered a measure of the hysteresis, increases with stress; the decrease at highest stresses is probably associated with plastic deformation. Preliminary work using only resistance measurements was done with an alloy containing 60.15 pct Cu and 39.79 pct Zn by weight. The results indicated higher values o-F M, in agreement with the known variation of M. with composition.' The effect of stress on M, (2°C per kg mm-') was of the same order of magnitude as that shown in Fig. 1 for composition E. An increase in hysteresis with stress was also found. The following experiment was made in order to investigate a partially transformed structure. A specimen of alloy E was cooled to — 85°C under a stress of 4.7 kg mm-'. Under these conditions, the martensitic transformation started but did not go to completion. The stress was then released and the specimen cooled to — 105°C. Fig. 2 shows the measured elongation c. The first change in the slope of the curve indicates the beginning of the transformation under stress. Removing the load at —85°C caused a decrease in length to the value corresponding to the elastic elongation of the parent phase resulting from the applied stress. Hence, the marten-site formed in the first part of the experiment apparently disappears completely and without hysteresis upon the release of the stress. The increase in length on further cooling indicates renewed formation of martensite. These conclusions are consistent with the concept of "thermoelastic" martensite," which has been confirmed by test." Acknowledgments The authors are greatly indebted to Professor M. Cohen for his advice and encouragement. They also thank F. Paxton for assistance. Thanks are due the American Brass Co. which supplied the alloys. References E. Kaminsky and G. V. Kurdjumov: Zhur. Tekhn. Fiziki SSSR (1936) 6, p. 984. A. B. Greninger and V. G. Mooradian: Trans. AIME (1938) 128, p. 337. "J. E. Reynolds, Jr. and M. B. Bever: Trans. AIME (1952) 194, p. 1065; Journal of Metals (October 1952). 'A. L. Titchener and M. B. Bever: Trai~s. AIME (1954) 200, p. 303; Journal of Metals (February 1954). " 3. A. Kulin, M. Cohen, and B. L. Averbach: Trans. AIME (1952) 194. p. 661; Journal of Metals (June 1952). "J. K. Pate1 and M. Cohen: Acta Metallurgica (1953) 1, p. 531. 'C. Crussard: Comptes Rendus (1953) 237, p. 1709. ' G. V. Kurdjumov: Zhur. Tekhn. Fiziki SSSR (1948) 18, p. 999. G. V. Kurdjumov and L. G. Khandros: Dokl Akad. Nouk. SSSR (1949) 66. p. 211.

Jan 1, 1957

By C. J. Johnson, C. J. Bise

Coal extraction by continuous miners (CM) is currently the most common underground method in the US industry and accounts for slightly more than two-thirds of the nation's deep mining production (National Coal Association, 1987). Even if longwall mining should become more commonplace, it can proceed only after ventilation and access entries have been driven by CMs. Since an area of concern continues to be the effects of the dust generated on the health of mineworkers, this paper discusses the relationship between the elemental compositon of mining-generated airborne dust sampled from the immediate ventilation return of a CM and laboratory-generated dust derived from channel samples taken from the mines. There are several potential contributions of this type of study to the coal mining industry. First, after more fundamental knowledge of the causes of Coal Worker's Pneumoconiosis (CWP) is learned, the laboratory-generated respirable dust could be used to identify a potentially hazardous coal seam. Also, this study could possibly aid in understanding the fundamental causes of CWP by producing mining-simulated samples of coal dust that could be used in epidemiological studies. Further. assuming that there is no difference in the elemental composition of a drill-core sample and a channel sample from the same location, a mining company could predict anew mine's respirable dust elemental composition in the immediate ventilation return by using exploratory drill core samples of the roof, coal, and floor rock to prepare the laboratory dust. Ventilation engineers could then use engineering design and control measures during premine planning to reduce the incidence and severity of CWP by better ventilating the potentially hazardous coal seam. If this proper planning prevented any future changes to the ventilation equipment and mine design. much time and money could be saved. Scope of work To investigate the variability of the chemical characteris tics of respirable dust, airborne dust samples from eight underground coal mines located in the eastern and midwestern United States were collected with eight-stage Sierra Model 298 Marple cascade impactors, as well as 25 channel samples of mined material. Each channel sample was removed from the middle of the coal face before mining occurred. Sampling of the mining-generated dust was conducted by Lee (1986) by sampling the entire working sections, primarily for characterization purposes, to obtain information on the locational variability of dust characteristics. Research performed for this study used the elemental analyses of the mining-generated dusts he sampled in the immediate ventilation returns of CMs. The procedure that was used to produce the laboratory generated respirable dust was based on the Hardgrove grindability test since it reflects the pulverizing characteristics of coal. This test was chosen for several additional reasons. First, it is repeatable and reproducible. A consistent amount of input energy is used as well as a specified size range of feed material to be crushed (the channel samples). Second, it is thought to generate secondary dust in a way similar to that of the crushing and grinding of the coal and rock as they pass through the arc-shaped cutting path of the CM's cutter head. The potential effect on dust generation by this secondary grinding mechanism may be at least as much as that produced by primary fragmentation, which is dust produced by the cutting action of the bit against the coal or rock (Roepke. 1984). Finally, the Hardgrove grindability test is well known and is used in the coal industry to guide mineral processing engineers in estimating the capacity of mills used to grind coal. One hypothesis of dust researchers in the Generic Technology Center for Respirable Dust is that the elemental as well as the physical characteristics of coal mine dust will make a difference in the incidence and severity of CWP. Coal mine dust is generated not only from coal, but also from any rock partings contained within the seam or any roof or floor material mined with the coal. Thus, coal mine dust may not have the same elemental characteristics as the coal being mined. Given that hypothesis, mixtures proportional to each thick¬ness mined of roof, coal, and floor rock derived from the channel samples of the face areas from which the respirable dusts were generated by the CMs were used to produce the feed material that was pulverized in the Hardgrove machine.

Jan 1, 1990

By J. Libaude

The Changement mine in French Guyana has, since 1983, processed gold ore by classical gravity methods. A large part of the gold is in the form of very small, free or enclosed grains. Recovery by gravity methods is thus limited (25-30%) and the present plant produces about 40,000 t annually of auriferous waste assaying about 6 git. The company therefore decided to retreat the waste by heap cyanidation. After laboratory tests which showed the feasibility of the method, a pilot test was set up on site. An experimental 3000 t heap, composed of waste of 0-8 mm grain size, was laid out by bucket loader on an impermeable pad covered by PVC and geotextile cloth. The upper layer of the heap was reworked by mechanical shovel to break up the area that had been compacted by the passage of the loader so as to restore adequate permeability.

Jan 1, 1998

By J. C. Martin

A theoretical analysis of steam stimulation is presented for single sands. The analysis includes the effect of steam production and most of the effects of heat conduction. The results show the effects of a number of important variables on the performance of an idealized well under steam stimulation. Calculated responses are presented which indicate the effects of steam production, amount of steam injected, water production, formation thickness and formation damage. Results indicate that steam production can cause large reductions in the heat contained in the formation. This effect can be eliminated by drawdown control. Water production reduces the amount of oil produced during stimulation. The optimum amount of steam to inject depends on economic factors as well as the well response. In many cases, the increased temperature resulting from stimulation reduces oil viscosity near the well sufficiently to overcome the effects of formation damage even if the damage is not removed during steam injection. Calculated responses for thin sands are more favorable than anticipated. INTRODUCTION Little has been published on the theory of steam stimulation'-' despite the interest it has created and the wide variation in well responses. The results of the present analysis provide an insight into steam stimulation, and the methods employed provide a foundation for future work. Analyses presented in Refs. 1 and 2 are very limited and apply to gravity drainage conditions. Ref. 3 contains an analysis similar to the one presented here. The idealized models used and the assumptions made in Ref. 3 are different from those used in this paper. The analysis assumes that after steam injection has heated a small portion of the volume within the radius of drainage of a single uniform sand, a shut-in soaking period is allowed before returning the well to production. The effects of gravity, capillarity, transient pressure and water-sensitive sands are neglected. The injection and soaking times are assumed short compared to the stimulated production time. The initial temperature is assumed uniform; thus, the results apply primarily to first-cycle stimulation. The effects of gas production other than steam are neglected, and the water-oil ratio during production is assumed constant. Steam stimulation involves the simultaneous variation of the temperature, pressure and saturations. General mathematical equations for these variations are complicated and can be very difficult to solve. Simplified equations based on idealized models are used to reduce the mathematics sufficiently to allow approximate solutions to be obtained. DISCUSSION INJECTION An idealized model for which heat conduction is neglected is used to represent the behavior of a well during steam injection. The mathematics for this model is presented in Appendix A which also contains an approximate solution for the behavior of the no-conduction model. SOAKING During soaking, the well is shut in. Only the temperature, pressure and saturation distributions at the end of soaking are needed in the analysis. During soaking the heat is considered to be conducted in a uniform medium from an initially uniformly-heated circular cylinder confined to the producing interval. At the end of soaking the saturations and pressures are assumed to correspond to the cold zone. Analysis of the heat flow during soaking is included in the next section. The radius of the heated cylinder is calculated from the following heat balance (for constant quality steam injection). At the end of the soaking period it is assumed that little or no free gas is present near the well, and that the soaking period has been sufficiently long that the steam zone has had time to expand and the steam has condensed. The condition where there is no soaking is considered in the next section. PRODUCTION In this section, an approximate method is presented for solving the equations of heat and fluid flow associated with the production of oil and water during steam stimulation. Where initial pressure drawdowns are sufficient to cause steam to flow into the wellbore, steam production is assumed to occur within a short initial adjustment period (Appendix B). The production practices followed soon after the well is returned to production can have a large influence on the amount of oil produced during the stimulation cycle. Under most conditions, there is a period of time in which some or all of the water in the heated zone is converted into steam and produced. This flashing of hot water is caused by the pressure in the heated zone dropping below

By H. C. Gatos, A. D. Kurtz

WITH the growing interest in the mechanism of self-diffusion of metals, the study of accurate and convenient methods for determining self-diffu-sion coefficients appears highly desirable. It was with this objective in mind that the present investigation was undertaken. Gatos and Azzam1 employed an autoradiographic technique for measuring self-diffusion coefficients of gold. This method involved sectioning of the specimen through the diffusion zone and recording the radioactivity directly on a photographic film. Because of the very short range of the emitted ß rays in gold, the activity recorded on the film was essentially the true surface activity. With proper choice of the sectioning angle, sufficient resolution could be obtained and the entire concentration-distance curve recorded in one measurement. For the boundary conditions of the experiment, where an infinitesimally thin layer of radioactive material diffuses in positive and negative directions into the end faces of a rod of infinite length, the solution of the diffusion equation is C/Cn = 1/v4pDt exp (-x2/4Dt) where C is the concentration of diffusing element (photographic density in this case), C,, is the constant (depending upon amount of radioactive material), x is the diffusion distance, D is the diffusion coefficient, and t is the time. Thus, by plotting the logarithm of the concentration vs the square of the diffusion distance, a straight line results and the slope contains the diffusion coefficient. In this manner, the self-diffusion coefficient of gold can be obtained as a function of temperature. In the present investigation the results reported by Gatos and Azzam1 have been verified, and the autoradiographic technique has been further developed and applied for the determination of the self-diffusion coefficient of gold at a number of temperatures. Furthermore, the energy of activation for the self-diffusion of gold has been conveniently determined. . Experimental Techniques Preparation of Specimens: The inert gold of high purity was received in the form of a rod from which cylinders were cut and machined to a diameter of 0.500 in. The specimens were annealed to a suitably large grain size and the faces were surface ground prior to the deposition of the radioactive layer. The radioactive isotope Au198 was chosen. It was produced in the Brookhaven pile by means of the reaction Au197 + n ? Au108. It decays by ß emission (0.96 mev) to Hg108 with the subsequent emission of a y ray (0.41 mev). 70Au 108 ? 80Hg 108 + -1e°. The half life of the Au108 is 2.7 days so that a strict time schedule had to be maintained in order to secure sufficient activity until the end of the experiments. For this reason, initial activities as high as 10,000 millicuries per gram were used. The gold arrived in the form of foil and was evaporated onto one face of each gold specimen cylinder to a thickness of about 100A. A sandwich-type specimen was formed by welding two such cylinders together. Evaporation of Gold: The gold was evaporated under vacuum from heated tantalum strips which were bent in such a way as to limit the solid angle through which the gold was allowed to vaporize, thus insuring a more efficient utilization of the gold. The specimens rested on flat brass rings which had an inner diameter of 0.475 in. The entire specimen-holding assembly could be manipulated from outside the vacuum system by means of a magnet which attracted a slug of soft iron attached to the assembly. By evaporating inert gold on glass slides under conditions identical to those employed for the radioactive gold, it was found that the thickness of the films was about 100A. Welding: The welding was performed by hot pressing in a stainless steel cylinder. The inside of the cylinder was threaded and fitted for two plugs. The specimens to be welded were placed in the middle of the cylinder and two pressing disks, one at each end, were inserted to avoid shearing stresses as the plugs were tightened. Mica disks were placed between the pressing disks and the specimens to prevent them from welding. The plugs were then tightened with a hand wrench and the entire unit was placed in an argon stream for about an hour to remove the oxygen. The unit was then inserted in the center of an argon atmosphere furnace maintained at about 700°C and left there for about an hour. Because of the difference in the temperature coefficient of expansion of the two metals, as the temperature rose. the pressure on the specimen-rollple increased and a weld resulted Welding was generally satisfactory under the conditions described.

Jan 1, 1955

Your Council has pleasure in submitting its report for the year ended 30th June, 1969. COUNCIL The following served as Office Bearers during the year under review: Messrs. R. C. J. Goode (President), J. K. E. Douglas and V. C. Robinson (Vice-Presidents), P. Lambooy (Immediate Past President) and D. G. Maxwell (Honorary Treasurer). Members of Council: Messrs R. J. Adamson, M. Barcza, F. P. Bath, Or J. M. Bereza, Or W. Bleloch, Messrs H. Britten, H. E. Cross, A. R. C. Fowler, A. E. Gilfillan, G. W. Holl, Prof D. D. Howat, Or J. P. Hugo, Messrs C. J. Irving, D. M. Jamieson, J. G. Kirchner, Prof J. de V. Lambrechts, Mr C. E. Mavrocordatos, Or J. T. McIntyre, Mr J. A. Nixon, Prof R. P. Plewman, Messrs J. F. Reid, H. Simon, C. G. Sowry, W. T. Stobart and P. W. J. van Rensburg. Ten Council meetings were held during the year with an average attendance of eighteen. FINANCE The annual accounts, which are attached to this report, show an excess of expenditure over income of R2,124, compared with an excess of income over expenditure of R4,218 last year. There are a number of reasons for this substantial change. There was a loss on realization of investments compared with a substantial surplus in recent years. The publication of the Journal once again resulted in a large loss in spite of a generous subvention of R2,000 from the Department of Cultural Affairs and other assistance amounting to R750. The revenue from sales of publications was down considerably. This is a figure which fluctuates considerably from year to year. The secretarial fee was considerably higher than previously. Due to pressure of work it was necessary for the Secretary to relinquish his duties with the Institute of Land Surveyors and, therefore, for this Institute to assume the additional staff costs involved. Finally there was unusual expenditure on the 75th Anniversary celebrations. There was a pleasing increase in revenue from subscriptions, due to the steadily increasing membership, but this was quite inadequate to cover the substantial increases in expenditure and losses in revenue mentioned above. The changes that have been effected in the arrangements for the publication of the Journal have been mentioned elsewhere in this report. It is expected that these will bring about substantial financial benefits in the long term. The MacArthur Forrest Memorial Fund shows an excess of income over expenditure of Rl00 and the total fund was therefore increased by this amount to R4,206. The balance sheet indicates that the market value of quoted shares and debentures increased during the year from R8,755 to R13,059, while the Accumulated Funds now amount to R22,567 as against R24,691 last year. MEMBERSHIP Fifty-six Members, fifty-five Associate Members, thirteen Graduates, twenty-two Affiliates and twenty- nine Students were elected during the year. Twelve Members were admitted to Life Membership; seven Associate Members were transferred to the grade of Member; four Graduates and two Students were transferred to the grade of Associate Member; five Students were transferred to the grade of Graduate; three Affiliates were transferred to the grade of Member; five Affiliates were transferred to the grade of Associate Member. Three Members, seven Associate Members, two Graduates, two Affiliates and six Students were removed from the Membership Roll in terms of By-Law 7.7.1. Your Council records with deep regret the death during the year of the following members: E. F. J. Bulman (Member), E. F. Chapin (Member), R. A. H. Flugge-de-Smidt (Hon. Life Member), T. S. Haldane (Member), J. M. Massie (Associate Member), S. W. Page (Student), S. R. Rabson (Member), M. Ramsay (Member), F. Wartenweiler (Hon. Life Member). A comparative statement of the membership as at 30th June, 1969, and the end of the previous year is shown in the following tabulation: 30th June 30th June 1968 1968 Honorary Life Members 12 10 Honorary Members 6 6 Life Members 176 189 Members 488 514 Associate Members 409 448 Graduates 60 68 Affiliates153 155 Students152 162 1,456 1,552 PAPERS AND JOURNAL The Institute is changing the format of the Journal from the present size to the standard international A4 size with effect from August, 1969. In making this change it will conform to more general practice today in regard to publications and this should effect certain economies and facilitate selling of advertising space. It has been arranged that in future the publication of the Journal, including advertisement sales, will be handled by Kelvin Publications (Pty.) Ltd. who are already doing this work satisfactorily for the Mechanical and Electrical Engineering Institutes. In making these changes the opportunity was taken of redesigning the outer cover of the Journal to incorporate the new coat of arms. The Presidential Address by Mr R. C. J. Goode and twelve papers which were published in the Journal during the year are listed below: August, 1968: Presidential Address: "The Institute-In retrospect and prospect." September, 1968: "Rotary mill liner practice in the South African gold mining industry," by J. H. French and O. E. Lissner.

By Larry D. Patts

Section 317(e) of the Federal Coal Mine Health & Safety Act of 1969 directed the Secretary of the Interior to prepare standards under which all working places in a mine shall be illuminated by permissible lighting while persons are working in such places. Section 317(e) further provides that such working places shall be illuminated within 18 months after such standards are promulgated. In accordance with this section of the Act, there was published in the Federal Register for December 91, 1970, a notice of proposed rulemaking which prescribed the illumination to be provided in the working places of underground coal mines. In light of written comments, suggestions, and objections to this proposed rulemaking, the proposedstandards were withdrawn and reproposed in the Federal Register for Wednesday, October 27,-19h. In light of further comments, suggestions, and objections, a public hearing was held on April 4, 1974, and standards were again reproposed and published in the Federal Register for April 1, 1976. Promulgation of the final lighting standards took place on October 1, 1976, which means that the underground coal mining industry must comply with face illumination requirements by April 1, 1978. As mentioned previously, the first proposed rulemaking for illumination of underground coal mines was published in the Federal Register on October 27. 1971. In early 1972, Consolidation Coal Company (Consol) and the United States Bureau of Mines agreed to a cooperative study of underground face illumination: Consol felt that expertise is this field would become increasingly important. Consol's initial efforts in illumination were aimed at investigating practical lighting systems for underground face equipment. We were concerned with installing unobtrusive lights which provided sufficient face illumination for safety, but at the same time were readily maintainable, electrically reliable, and physically sheltered from damage. We believe that our initial lighting systems did provide sufficient face lighting for safety, but because only prototype components were available for field testing, the resultant poor system reliability and maintainability necessitated drastic improvement before face lighting could become practical. Final Lighting Standards Deem Early Lighting Installations Out Of Compliance On April 1, 1976, the Federal Register contained the final version of the illumination standards (as they were later promulgated in October). When these illumination regulations and measurement techniques were defined and measuring instruments were available, Consol checked their lighting systems underground and determined that the systems were not in compliance with these final illumination standards. More Lighting Hardware Added In An Attempt To Achieve Compliance. After determining that all of our face lighting systems were not in compliance, we began adding additional lighting hardware in order to meet compliance with published regulations. Unfortunately, to date, we have not been able to meet compliance with practical lighting systems. We have determined from our field installations that the required additional lighting hardware, (to meet compliance) with its increased vulnerability and decreased reliability, renders the lighting systems impractical, if not impossible, to reasonably maintain. Our attempts to provide 0.06 footlamberts have also produced adverse operator reaction to the glare and to illumination systems in general. BCOA Members Ask MESA To Demonstrate Practicability Of Compliance With Regulations Industry concern about meeting the impending lighting regulations was mounting, and in May of 1976 a meeting between MESA and BCOA members was held to discuss lighting compliance problems. At this meeting, BCOA offered to work cooperatively with MESA in testing the practicability of various lighting systems mounted on underground mining equipment. Field tests were to be conducted by United States Steel Corporation, American Electric Power Service Corporation, and Consolidation Coal Company. The purpose of this underground field testing was to develop capability to provide adequate face illumination in a safe, workable manner which would not detract from efficiency of operation. BCOA members involved in this cooperative study were to submit necessary machine drawings, sketches, etc. to MESA in order that MESA could perform a "black-box" study and specify the type and location of luminaires to be installed on the test machines. MESA was confident that they could specify lighting systems that would be in compliance and would be practical so as not to detract from efficiency of operation. Consol was first to install lighting hardware under the BCOA/MESA cooperative agreement. As per MESA specifications, Control Products HgV luminaires were installed on a Joy 2BT-2H boring machine at the Williams Mine of Northern West Virginia Region. As of mid-October, 1976, Consol had approximately eight weeks operating experience with the lighting system on this boring machine underground and had drawn the following conclusions: The lighting system installed at Williams Mine (1) does not meet compliance with lighting standards as originally proposed by MESA, (2) does not provide illumination in a safe workable manner, and (3) will detract from efficiency of the mining operation due to operational delays. Although Consol has rearranged lights on this boring machine in an attempt to reduce operator objections, a practical lighting system which is "in compliance" has not been arrived at as of this writing.

Jan 1, 1979

By M. I. Sherman, J. D. H. Strickland

USE of a hydrometallurgical approach to the oxidation of sulfide ores and extraction of metals therefrom may have advantages over the more common smelting techniques when a low grade deposit is difficult to concentrate or the subsequent separation of metals, coexisting in the ore, is laborious by any known smelting operation. For economic reasons, the most promising oxidants are either atmospheric oxygen or electric power. The use of oxygen, or air under pressure, has recently been revised. Pyrrhotite has been converted to iron oxide and elementary sulfur' and a variety of sulfides have been treated by Forward and co-workers.2-4 Generally sulfate is the end form of the sulfur but with galena in an acid medium, elementary sulfur can be formed." For economic reasons chlorine and ferric iron salts are about the only possible alternatives to the atmosphere as oxidizing agents for base metal sulfides. If aqueous solutions of chlorine or ferric iron are employed, the reduction products can be oxidized electrolytically in situ and used again, thus acting as catalysts for electric power as oxidant. The use of ferric salts for this purpose is established hydrometallurgical practicea but, although chlorine gas has been employed in the dry state at an elevated temperature, its use in aqueous solution at or near room temperature has not found favor. The reaction of chlorine water with the soluble sulfide ion has been studied by several workers,7-9 and both sulfate and elemental sulfur are found as end products, the latter being favored by the presence of a low concentration of oxidant relative to that of sulfide in solutions of about pH 9 to 10. Of direct bearing on the work in hand are an early American patent" and a recent Austrian patent." The former advocates stirring powdered ore with an aqueous solution of ferric chloride chlorine oxides and chlorine. In the latter it is claimed that both metal and sulfur can be obtained by electrolysis, in a diaphragm cell, of a metal ore slurry in brine. Details in these patents are scant and no data or explanation is given for the mechanism of the reaction which, in the Austrian work, is attributed to the (unlikely) action of nascent chlorine at the anode surface. No mention is made of possible differences in behaviour between various ores. Apparatus A complication encountered when working with chlorine water is that a serious loss of chlorine occurs by gas partitioning unless an enclosed system is used and any air space in the apparatus is kept very small and constant. Arrangements were made, therefore, to take out samples for analysis without letting air into the system to replace the liquid removed. For convenience in studying a heterogeneous reaction the apparatus was so designed that a reproducible controlled stirring rate could be maintained and the ratio of surface area of ore to volume of solution was approximately constant throughout any experiment. The apparatus used is shown in Fig. 1. The ground ore was placed in the horizontal cylindrical vessel, A, of about 1 liter capacity, heated by a constant temperature circulating bath pumping water through the concentric jacket, B. By adding chro-mate to this water, an ultraviolet radiation filter effectively surrounded the reaction vessel, greatly reducing any possible photochemical decomposition of chlorine solutions. Stirring was effected by glass paddles, C, attached by an axle to a magnet which was rotated by another powerful Alnico magnet, D, outside the glass end, this magnet being itself rotated by an electric motor electronically controlled to constant speed. Speed could be varied from about 150 to 900 rpm and was measured and held to within 1 pct of a given value. The end of the reaction vessel remote from the stirring magnet was closed by another one-ended glass cylinder, E, connected by thin polyethylene bellows, F, clamped by screw clamps and watertight rubber gaskets to the main vessel. Through E, a glass electrode and calomel electrode projected into the solution and a hypodermic syringe pierced a small bung and allowed acid or alkaline to be added to maintain a constant pH. By pushing the fully extended bellows until the two cylinders touched, from 50 to 100 ml of solution could be forced out through a sintered disk into the three-way tap system, G, either to waste (for flushing purposes) or up into a 10 ml burette where the solution could subsequently be measured out for analysis. The ore samples were introduced at H, the tube being stoppered by a thermometer of —1 to +52ºC range, graduated to 0.1°C intervals. To prevent ore from being ground in the end bearings of the stirrer these bearings were pro-

Jan 1, 1958

By H. R. Froning, R. O. Leach

In wertability alteration flooding, a chemical agent is rnoved through a reservoir by the flood water to increase oil recovery by decreasing the degree of wetting of the rock by the oil. Substantial amounts of the chemical may be lost during movement through the reservoir. The extent of the loss, and therefore the economics of the process, depends in some cases on factors which are difficult to reproduce in the laboratory. Therefore, a short-duration, low-cost field test method is needed to permit evaluation of chemical requirements under actual field conditions. This paper describes a small scale rest conducted at a single well for measuring chemical requirements, thereby giving a more reliable evaluation of this important factor in the applicability and economics of the process. In the rest a small wafer slug containing the chemical agent and a nonadsorbed tracer is displaced into the reservoir by a known volume of wafer. The well is then placed on producrion. Chemical loss per barrel of pore volume contacted is calculated from [he fractional recoveries of the agent tested and the nonadsorbed tracer. The method has been used to determine within the actual reservoirs the chemical requirements for both a sandstone and a dolomite reservoir. Several chemical agents are potentially available for wettability alteration flooding, although none is universally applicable. For some applications of the method, chemical costs per barrel of additional oil recovered can be substantially less than one dollar. INTRODUCTION Wettability alteration flooding provides a means of increasing oil recovery from reservoirs by decreasing the degree of wetting of the rock by the oil and increasing the displacement efficiency of the flood water. Earlier studies demonstrated a relationship between oil recovery during waterflooding and the degree of wetting of a rock surface by an oil. The application of wettability alteration flooding to the Harrisburg field of Nebraska provided a field test' of this recovery process. Subsequently, additional laboratory and field tests have developed additional procedures for evaluating wettability alteration flooding, and have indicated where the process may be applicable. Applicability of this process to specific reservoirs is determined by a progression of tests to determine sus- ceptibility of the reservoir to alteration of its wettability, to indicate the degree of recovery improvement and to estimate the amount of chemical required to process the reservoir. The economics of applying improved oil recovery processes depends not only upon the degree of improvement in oil recovery achievable by the process but also upon the process costs and the timing of the income and the investment. Emphasis in this paper is on the expenditure aspects of the process. The work reported in this paper indicates that the chemical investments required for wettability alteration flooding are substantial. For evaluating the economics of a potential flooding application it is imperative that a sound estimate of the chemical requirements be made for the reservoir. Generally, true reservoir conditions are not adequately simulated in laboratory chemical propagation tests. Because of wide well spacings, many years might be required to obtain chemical propagation data from conventional pilots or inter-well tests. Consequently, n short-duration, low-cost method is needed to determine chemical requirements in the field. The potential applicability of wettability alteration flooding is discussed, as well as the economics of wettability alteration with respect to the inherent and imposed restrictions on the timing of income and investments. DETERMINATION OF CHEMICAL REQUIREMENTS In the process of moving a chemical bank through reservoir rock, some of the chemical agent lags too far behind the flood front to be effective or is otherwise lost to the reservoir system. The extent to which these losses occur overshadows the reductions in chemical concentration due to diffusion and to mixing with the reservoir fluids. Experience indicates that almost without exception, chemicals which induce a wetting change undergo either sorption reactions or chemical reactions with mineral constituents of the pore surfaces. Other reactions may occur between the added chemical and the reservoir oil and water. Even in limiting consideration to reactions of the relatively inexpensive inorganic salts, bases and acids, the reactions may be exceedingly complex. Reservoir pore surfaces consist of more than silica in sandstone reservoirs, and more than calcite or dolomite in carbonate reservoirs. Many mineral species are present, each exhibiting specific tendencies to react with an injected chemical. The reactions which occur can consume enough of the agent to have an important effect on economics. These reactions can cause a change in pH of the chemical bank, or may remove some of the active chemical by precipitation, adsorption or reaction to form a new chemical which may or may not be effective in

By Olaf Hondrum

CHURN drilling equipment at Bagdad consists of two Bucyrus Erie 27-T model drills and one 22-T drill with gasoline engines. The drilling tools weigh approximately 1600 lb. The holes are drilled with 7-in. bits which are heated in an oil burning furnace and sharpened in an electrically powered bit sharpener. A service truck brings the bits to the drills and takes the dull bits back to the shop. The bits weigh around 250 lb. when new and are handled by drill crews without any mechanical handling device. When the management decided to start open- pit operations in 1945 the ruggedness of the terrain was one factor that led to the purchase of the small 22-T model drill. The orebody lies on both sides of Copper Creek Canyon. The sides of the canyon are steep and are cut by numerous narrow branch gullies. The drill had to travel over roads that were narrow and steep. It proved adequate for drilling the capping formation and is still in use.

Jan 1, 1950

All companies and individuals who embark on a business venture incur costs. For the most part these costs can be conveniently categorised, firstly into capital and operating costs and then into the various elements of these major cost centres. Although the nature of expenditure naturally differs as between industries certain broad cost types are characteristic of all heavy industries. The capital costs of gaining access to an orebody, equipping a new mine, building a treatment plant and perhaps refining and smelting facilities are likened to capital expenditure by a manufacturer on new plant and equipment required for an expansion or new product innovation. Similarly the day to day costs of labour and materials employed in the running of a mining operation can be likened to similar costs in other heavy industry. There is another group of costs that are not so easily categorised. Most industries incur unavoidable costs that do not form part of the direct or capital costs of physical production. These costs include taxes and other payments to government and research expenditure. The mining industry has a particularly heavy burden of non-production costs and it is the objective here to highlight and analyse a few of them. These costs are exploration expenditure, infrastructure expenditure, royalties and income tax. This is not to suggest that these are the only major non-production costs that the mining industry needs to bear. The industry may incur significant costs for environment protection, land tenure maintenance and on liaising with Government on a wide variety of matters. It is for reasons of subject manageability that the list is confined to the four costs mentioned. For the mining industry, exploration and infrastructure expenditure, royalties and income tax have several distinguishing features. Firstly, they are generally proportionately large costs. Exploration and infrastructure expenditure for example, could well exceed mine development costs. Depending on how royalties are levied, they may be very high in relation to profits earned and many in fact exceed profits. Secondly, except for income tax, they are costs that are generally not payable by other industries. Exploration expenditure is in some respects comparable with industrial research and development. However, because mines are wasting assets exploration is necessary for survival, and it is likely to form a larger part of costs per dollar of profits or sales than R&D. It can therefore be said that exploration is of far more significance to the miner than R&D is to the typical industrial company. Furthermore, unlike exploration expenditure, R&D results are transportable and patents permitting, transferable. The infrastructure burden that miners must carry is unique to the industry. Whereas urbanised industry is not usually liable for the capital cost of power, water, transport, townships and community facilities, the miner must finance these projects himself. Neither the payment of royalties, nor of course, income tax are unique to the mining industry. Royalties are more properly described as a unique charge on some primary industries but however described, they are a charge that manufacturing and service industries do not have to pay. Whilst we all pay income tax, the mining industry's uniqueness derives from the special problem that the industry has in arriving at annual net income in view of the capital intensity of the industry, partly due to expenditure on infrastructure, the indefinite life of capital equipment and its unpredictable residual value, funds expended on exploration and long lead times associated with a mine development. The third distinguishing feature of these costs is that, with the possible exception of exploration expenditure, they are heavily influenced by the policies and actions of Government. As such the industry has a responsibility to promote the industry by emphasising its potential to improve national product including exports and by requesting that the Government provide for the industry in its long term planning and seek its co-operation to prevent the industry incurring inequitable penalties relative to other industries.Finally, exploration and infrastructure expenditure are significant because they are front-end charges on a project. When due allowance is given the time value of money and the deferment of income from a mining development to which they relate, exploration and infrastructure expenditure assume greater importance than the raw money expenditure suggests. For all these reasons, exploration and infrastructure expenditure, mining royalties and mining income tax are worthy of detailed study. This is the purpose of this paper. The objective is to examine the determinants that give rise to them and provide some thoughts on their impact on on-going and potential mining developments and, where relevant, to make some suggestions as to how they may be minimised.

Jan 1, 1977

By C. H. Bowen

NDUSTRY needs a generally applicable means of defining average grain size and grain size distribution. Students of sediments hade explored this field, employing methods that might also prove useful in engineering problems. Before attempting to solve specific problems it is well to review the derivation of commonly used grade scales and the reasons for their selection. This aspect of the problem seems largely to have been lost, and a review of basic factors may suggest causes for failures in using size analysis data. Three facts are implicit in selection of a grade scale: 1) Most particulate mixtures are continuous distributions of sizes, and any grade scale that may be employed is an arbitrary means of visualizing that distribution. 2) For purely descriptive purposes, any grade scale, regardless of the rationality of the class intervals, will be satisfactory if it is accepted by a sufficient number of workers. 3) For analytical purposes, class intervals must be small enough to define the continuous distribution accurately. Further, where statistical studies are involved, a fixed relationship should exist between classes or grades. An argument in favor of geometrically related size grades lies in the fact that most particulate mixtures contain such a wide range of sizes that use of an arithmetic diameter scale is practically impossible. Udden, who recognized this fact in 1898,' proposed one of the first grade scales based on a regular geometrical interval. Udden used 1 mm as his basic diameter and a ratio of 2 (or 1/2) between classes. In 1922 Wentworth2 re-examined Udden's grade scale, retaining the same class interval and basic diameter, but extending the scale in both directions and renaming the classes. In 1930 (Ref. 3, p. 82) the American Society of Testing Materials proposed what is now known as the U.S. Standard fine sieve series, also based on the 1 mm diam, with a v2 ratio between sieves. This, then, is a one fourth Udden-Wentworth series in the sizes below the 4 mesh sieve. The U.S. Standard coarse sieve departs from the 1 mm base and uses inches; hence it is not a direct continuation of the fine series. The U.S. Standard series would thus seem to possess all the attributes of a good grade scale, which it is. It has a large number of classes (sieves). in fact too many for practical use in its entirety. This v2 subdivision of the Wentworth grades has led to the common use of two v2 sieve series, the half-Wentworth and the engineers' series. Geologists and sedimentologists favor the half Wentworth, or 18, 25, 35, 45 sieves, etc., whereas the engineers, preferring round numbers, utilize the other half of the U.S. Standard grade scale in the 16, 20, 30, 40 sieves, etc. The fixed geometrical ratio between classes is an advantage in statistical analysis, but the unequal classes cause some complications in calculations. This is especially true when moment measures are used. It was to simplify these calculations that Krumbein in 1934 devised the phi scale. Phi is defined as being equal to —log2 of the diameter in millimeters. Selection of logarithms to the base 2 relate the phi scale directly to the Wentworth grade scale in such a manner that the whole or fractional diameter values 2, 1, 1/2, 1/4 mm, etc., become rational whole numbers, —1, 0,1, 2, etc. Since this is an arithmetic rather than geometric series, calculations are facilitated. When the logarithm is multiplied by —1 the phi values below 1 mm become positive, those coarser than 1 mm negative. Because of its relationship to the Wentworth grade scale (and in turn to the U.S. Standard fine sieve series) it is not necessary to use the transformation equation to calculate the phi value for each individual sieve; this can be done graphically as shown in Fig. 1. It should be noted that this graph may be extended in either direction to include the range of sizes most commonly used by the individual worker. Application to Statistical Analysis Any attempt at systematically relating size analysis data to properties involves a statistical study whether it is recognized as such or not. Since this is true it would seem more logical to use measures and devices related to the general body of statistical theory. Several methods are available for studying particulate mixtures. One of the most commonly employed, and also the most often misused, is the histogram or block diagram. If its limitations are recognized and provided for, the histogram is a very useful tool. According to conventional practice, the bars of equal width are plotted and the values noted in terms of diameters, when in point of fact, log diameter is implied by such notation. Further, the histogram is sensitive to choice of grade scale and size of class interval, either of which may color the result. Grade scales whose classes are not related by fixed intervals are particularly difficult. Another basic weakness of the histogram is that it pictures a continuous distribution as a series of discrete grades.

Jan 1, 1957

"Boston’s Central Artery passes beneath nine active railroad tracks carrying commuter and Amtrak trains into South Station, the City's busiest rail terminal. These railroad operations had to remain in service throughout the construction of the “Big Dig”. The solution was to jack the tunnel under the tracks, rather than build the tunnel by more conventional means.Jacking of utility pipes is fairly common, but jacking boxes as big as a highway had never been done in North America, until the Big Dig came along.Soil freezing was decided upon as a means of stabilizing the soil under the tracks. This allows the tunnels to be jacked into place without soil caving when the soil is excavated at the leading face of the advancing tunnel boxes. Hundreds of pipes were driven into the ground between the tracks. Inside the steel pipes were smaller plastic pipes. A freezing plant near the railroad tracks pumped refrigerated brine into the inner plastic pipes. The brine ran out the bottom of the inner pipe, flowed up to the top inside the outer pipe, and was re-circulated to the freezing plant and back again into the ground. Over several weeks the circulating brine drew the heat out of the soil little by little, and froze the ground outward from the pipes. Soil FreezingThe plan required the construction of jacking pits in which the tunnel boxes could be constructed, and from which they could be jacked forward to their final locations, as described in the accompanying article. Once the jacking pit was built and the tunnel box constructed inside the pit, crews made an opening in the headwall of the pit to allow passage of the tunnel sections, excavated about three feet of the frozen soil, pushed the tunnel box ahead, and repeated the procedure again and again."

Jan 1, 2003

By David C. Russell

The Department of the Interior used to be a quiet, noncontroversial, almost boring agency. It, after all is the fifth oldest of the Departments, and as an old line Federal agency it has studiously performed its preservation and resource management functions in a caretaker mode--though some would say more "undertaker" than "caretaker"--locking up the body and soul of America piece-by-piece. Yes, quiet, serene. That is until Jim Watt showed up. And we have all seen that version of Mt. Vesuvius which resulted--only it was the environmentalists who blew their tops. Ronald Reagan chose Jim Watt as Secretary of this fine old agency to prove that one-third of our Nation's land and over a billion acres on the Outer Continental Shelf can work for this Nation. At the foundation of President Reagan's charge to Secretary Watt is a belief in the tenets of the free enterprise system, and in the individual freedoms upon which this country was founded. There are those who don't share this belief in democracy and free enterprise, and those who believe this 205 year experiment called the United States of America will fail. Nikita Krushchev said "we will bury you"--obviously he didn't agree with our system. An Italian sociologist, Franco Ferrorotti, said bureaucratic stagnation will kill capitalism. Certainly we have all felt the ravages of bloated bureaucracies. Perhaps one indicator in the United States is the Federal Register, that daily compilation of Government's largesse. In 1970, 20,000 pages of the Federal Register were published. A decade later, in 1980, that volume had quadrupled to 80,000 pages. The Federal bureaucracy can stagnate from excessive budgets as well. The Interior Department spent $60 million on administering Federal coal leasing in 1981. That's nearly two bits a ton for every ton of coal leased in 1981. You wouldn't stay in business very long if your administrative overhead on inventory was that outrageous. But the pessimism of our critics is apparent from more than red tape and bloated budgets. For decades America has been fasting--consuming too little of America's wealth of minerals, subsisting instead on a diet heavily reliant upon mid-east oil, with little emphasis or concern for inventorying and developing domestic energy and mineral resources. Economics--yes. But short-term, short-sighted economics. Excessively dependent upon foreign imports, of oil, cobalt, chrome and other strategic minerals, the U.S. measures its time before another embargo--or fallen Shah, or Soviet manipulation, or Saudi shift, or, as we witnessed in Egypt, assassination--an untimely loss to mankind and efforts to bring peace to the troubled mid-east. These disruptions, in addition to their tragic human tolls, impair the free world's security. Huge chunks of the United States have been locked away in dozens of single land use categories in the name of conservation, with only the foggiest idea of what resources might be denied the American people-and this at a time of unacceptable levels of energy and strategic mineral imports. More than half and perhaps two-thirds of all Government-owned lands are totally withdrawn from or severely restricted to development under the mining and leasing laws. We must continue to rid Government of the overly zealous restraints which have been keeping us from drawing upon that which can help restore our economy and national security. When we assumed responsibility, the United States was dependent on foreign sources for about 40 percent of its oil. In 1981, our oil import bill was approximately $83 billion--nearly 17 times what it was in 1972. Our reliance on foreign sources for essential minerals is even more disturbing. We must look to other countries--some unfriendly, some unstable--for 22 of 36 strategically critical minerals. Yet the energy resources on federal lands which are owned by the American people could meet our needs for centuries if properly managed. Eighty-five percent of the crude oil yet to be discovered in America is likely to come from public lands, as will 40 percent of the natural gas, 35 percent of the coal, 80 percent of the oil shale, nearly all of the tar sands, and substantial portions of uranium and geothermal energy. Our vast hardrock-mineral wealth includes untapped deposits of essential elements we now import, such as chromium, copper, platinum, and cobalt. The obvious question is, if these abundant resources can help to revitalize our economic strength and to preserve our national security, why aren't we using them to better advantage? To a large extent, the answer can be found in past decisions to restrict public access to the federal estate, thus deferring to us or our successors the tough decisions that flow from Congress' mandate to provide for environmentally responsible development of America's energy and mineral treasures. Here is the legacy this Administration inherited: In January 1981, 7 years after the onset of the Mideast oil embargo: ---Less than 15 percent of federal onshore lands were under lease for oil and gas development; ---No oil and gas leases had been issued in Alaska for 15 years;

Jan 1, 1982

By R. Maddi, I. R. Kramer

The delay time for the initiation of slip was studied in single crystals of a brass, aluminum, and ß brass. A delay time for slip was found in ß brass when the specimens were tested below room temperature; however, one was not found for a brass or aluminum. A general theory for the existence of the brittle transition temperature is proposed. A LTHOUGH a considerable amount of effort has A been devoted to the study of the deformation and fracture of single crystals, the vast majority of the work was concerned with static tests or with tests carried out at relatively slow strain rates. The necessity for the study of a possible incubation time for slip becomes apparent when the various theories which have been postulated for the elucidation of the mechanism of slip are considered. The existence of an incubation period would strongly indicate that slip occurs by a process of nucleation and growth; whereas the absence of an incubation would present rather convincing evidence that slip occurs by a cataclysmic process. In addition to shedding light on the process of plastic deformation, an understanding of the early stages of plastic deformation may be helpful in finding an explanation for the brittle behavior of the body-centered metals under certain conditions. It is well known that these metals, such as iron, molybdenum, tungsten, ß brass, etc., will become brittle as the test temperature is lowered or the strain rate is increased. In spite of the fact that this subject has received considerable attention since the turn of the century, no adequate theory exists today which explains the phenomenon. In the ductile temperature range the metal breaks with a "fibrous" fracture after considerable slip has taken place. In the brittle temperature range the metal fails essentially by cleavage on the {loo) planes: however, some plastic flow is always present even in the most brittle fractures. It is possible, as will be explained later, to ascribe this brittle behavior to the length of the incubation time for slip. Clark and Wood,' using polycrystalline metals, showed that a delay time existed for the yield point in mild steels and that the delay time depended upon the applied stress. These authors stated that the delay time was observed only with those materials for which the stress-strain curve showed a definite yield point. In their work only the mild steel specimens exhibited a delay time at the yield point, while the other materials studied-type 302 stainless steel, SAE 4130 normalized steel, SAE 4130 quenched and tempered steel, 24s-T aluminum—did not show a delay time. These authors' in an extension of their work, found that as the temperature was lowered the delay time was increased. It is, then, the purpose of this investigation to measure or to set an upper limit on the delay time for plastic flow in single crystals of a brass, aluminum, and ß brass. The delay time will also be studied as a function of temperature. The incubation time may be measured by determining the length of time the specimen will support a stress without slip occurring when the stress is greater than the static critical resolved shear stress. In order to accomplish these measurements use was made of a pendulum which was so designed that a single crystal specimen could be placed at various positions along its length. When a pendulum is struck by another pendulum of the same length, an elastic wave is transmitted down the bar with the velocity of sound in the material and is reflected from the far end of the bar. The reflected wave then travels back and unloads the stress until it reaches the impacted end of the bar and the two pendulums separate. The length of time that the specimen is subjected to the stress is equal to the time it takes for the stress wave to travel twice the distance from the specimen to the end of the bar. The stress applied to the specimen is governed by the velocity of impact of the two pendulums. The stress at which the specimen first suffers plastic: deformation was determined in these experiments by examining the specimen under the microscope for the first appearance of slip lines and by measuring the residual strain after each impact. In these experiments the critical resolved shear stress was approached from the low stress side and no attempt was made to determine it exactly. To determine it exactly, it would have been necessary to find that stress at which slip would have just started. However, since the stresses were increased in rather small increments the critical stress is believed to be approached rather closely, as will be seen from the experimental results. The critical stress will be taken as the highest stress obtained before the onset of plastic deformation. The stresses

Jan 1, 1953

By J. Quigley

As the outcrop mines in the West developed into underground operations, systems of ground support were gradually evolved. In the early coal mines there was little need for support except near the dirt line in portals, where stone masonry was common. Where the top was shaley or broken, native pine props with light cross bars and legs furnished enough support even in Utah's 25-ft coal seams. As depth of workings increased. roofs and backs of the same general nature as those near the surface became more and more unstable and required more and more support. Some coal airways show this tendency very clearly. From the surface down the same type of roof shows deterioration which an experienced eye can translate into a measure of depth under surface rather than change in rock characteristics. Rock bolts, developed by various companies and by the U. S. Bureau of Mines, have become an effective substitute for timber in sections of some metal and nonmetal mines formerly requiring escessive timber support, and further use of war surplus landing mats, chain link fencing, and a new punched channel developed by one of the steel companies has enabled other mines to operate deposits where costs of timber and lack of clearance for timber support would have prohibited mining. The block caving mines have made extensive use of reinforced concrete underground to achieve similar ends under difficult conditions. Steel sets are standard in many Bureau of Reclamation projects, although these are usually covered in with concrete to make the permanent structures the Bureau's reclamation projects require. But the use of steel in mining operations is limited and has been confined principally to the iron ore mines of Michigan, Wisconsin, and Minnesota. Some mines have installed used rail as posts, caps, and crossbars, but a rail section is not suited for load carrying, and used rails are generally brittle. having a tendency to fail without warning when overloaded. European mines were the first to reach the size of worked out areas and depths of cover resulting in major roof problems. The Europeans resorted to pack walls and masonry walls, in conjunction with timber arched sets. rail arches, and combination timber and rail and steel arches. The give in these pack walls and wooden blocking was supplemented by a hinge in the center of the arch. This design is called an articulated arch Through various refinements of this principle of the support giving graduallv with the load. Toussaint-Heintzmann developed the yielding or sliding arches, in which yield is accomplished by friction in the overlapping joints of the arch. This type has gained widespread acceptance in the Ruhr and Lorraine Basin and is being manufactured by Bethlehem Steel for sale in this country. In North America the anthracite mines in Pennsylvania, followed by certain iron ore mines in upper Michigan and Canada, were the first to employ these arches to any extent. The practice was later adopted by Kennecott at Ruth, Nev., and by others. Despite high initial cost, the use of these arches is growing in many parts of the country because of their suitability in heavy ground. In its present form of manufacture the yield-able arch consists of open U-shaped rolled section with heavy beads on the edge. The open edge of the U is placed toward the wall. The section nests in another section of the same dimensions, and an arch can be built up from rolled radii and tangents of various weights and lengths. Sections are fastened together by U bolts and saddles. The lap on the joint varies from 12 to 24 in., and ordinarily the bolts are tightened with a 1-in. drive air wrench. The arches are spaced with channel struts held by J bolts and saddles. Sections can also be obtained that are composed of various combinations of radii and tangents and true circles. The joints can be placed to bear against anticipated loads and asymmetrical loads imposed by dipping strata. In the arches now being manufactured clearance widths up to 19 ft are obtainable in weights of sections from 9 to 30 lb per ft. The circular cross sections are available in the same weights ranging from 8 to 16 ft diam. At present most of the arches sold are supplied only in carload lots. It is hoped that demand will grow so that distributors can stock various weights and sections to give small operators a chance to try this new type of rock support under their own particular conditions. Several excellent papers have discussed the properties of various sections now manufactured, the dimensions of the sets obtainable, and their application under widely differing conditions. The present article will describe the methods and results of a special use of the arches at Kaiser Steel mine No. 3. Sunnyside, Utah. Problem at Mine No. 3 : In 1953 Kaiser Steel Corp. laid out Sunnyside mine No. 3 to recover coking coal left by the previous operator, Utah Fuel Co.. below workings that had been abandoned in 1928. Two seams had been worked, the upper and lower, separated by 30 to 42 ft of rock. Approximately 10 million tons of coal had been extracted from this area some 3000 ft down the itch from the outcrop to a 1500-ft depth of cover. The mine had been opened by slopes in both upper and lower seams. Sometime in the late 1920's the lower slope

Jan 1, 1960