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By R. L. Bell, T. G. Langdon

A technique was developed- for determining the grain strain, and hence the grain boundary sliding contribution, occurring during the high- temperature creep of a magnesium alloy, from the distortion of a grid photographically printed on the specimen surface. The results were compared with those obtained from measurements of grain shape, both at the surface and interrwlly, and it was concluded that the grain shape technique may substantially underestimate the grain strain and overestimate the sliding contribution due to the tendency for migration to spheroidize the grains. ALTHOUGH a considerable volume of work has been published on the role of grain boundary sliding in high-temperature creep, many of the estimates of Egb (the contribution of grain boundary sliding to the total strain) have been in error due to the use of incorrect formulas or inadequate averaging procedures.' One of the most easy and convenient measurements from which to compute Egb is that of v, the step normal to the surface where a grain boundary is incident. Unfortunately, this parameter is also the one associated with the treatest number of pitfalls. Values of v have been used to calculate Egb from the equation: egb =knrVr [1] where k is a geometrical averaging factor, n is the number of grains per unit length before deformation, v is the average value of v, and the subscript ,r denotes the procedure of averaging along a number of randomly directed lines. If the dependence of sliding on stress were assumed, it would be possible, in principle, to calculate k from the known distribution of angles between boundaries and the surface. This in itself is difficult because the distribution depends on the history of the surface,' but the problem is even further complicated by the fact that v depends on other factors such as the unbalanced pressure from subsurface grains.3 However, the great simplicity of the measurement procedure for v makes it highly desirable that this problem of k determination should be overcome. In the present experiments, this was achieved by the use of an indirect empirical method in which the grain strain, eg, at the surface was determined by the use of a photographically printed grid. The assumption here is that the total strain, et, is simply the sum of that due to grain boundary sliding, egb, and that due to slip or other processes within the grains, eg. SO that: Et = Eg + Egb [2] Thus k is given by: In practice, it is customary to indicate the importance of sliding by expressing it as a percentage of the total creep strain; this quantity is termed y (= 100Egb/Et). The determination of Eg from a printed grid within the grains avoids the difficulties due to boundary migration which should be considered when the grain strain is calculated from measurements of the average grain shape before and after deformation. As first pointed out by Rachinger,4,5 however, this latter technique has the particular advantage that it can also be applied in the interior of a polycrystal. Recently, several workers have produced evidence on a variety of materials6-'' to support the observation, first made by Rachinger on aluminum,4,5 that 7 can be very high, 70 to 100 pct, in the interior, even when the surface value, determined from boundary offsets, is very much lower.10'11 Although there have been criticisms both of the shortcomings of the grain shape technique'' and of the different procedures used to determine y at the surface,' it seemed important to check whether measurements of sliding by grain shape gave values of y which were truly representative of the material. In the present experiments, grain shape measurements were therefore made both at the surface and in the interior for comparison with one another and with the independent measurements of grain strain using the surface grid technique. EXPERIMENTAL TECHNIQUES The material used in this investigation was Magnox AL80, a Mg-0.78 wt pct A1 alloy supplied by Magnesium Elektron Ltd., Manchester. Tensile specimens, about 7 cm in length, were prepared from a 1.27-cm-diam rod, with two parallel longitudinal flat faces each approximately 3 cm in length. The specimens were annealed for 2 hr in an oxygen-free capsule, at temperatures in the range 430° to 540°C, to give varying grain sizes, and, prior to testing, the grain size of each was carefully determined using the linear intercept method. This revealed that the grains were elongated -0.5 to 5 pct in the longitudinal direction. Testing was carried out in Dennison Model T47E machines under constant load at temperatures in the range 150" to 300°C. At temperatures of 200°C and below, tests were conducted in air with the polished flat faces coated with a thin film of silicone oil to prevent oxidation; at higher temperatures, an argon atmosphere was used. To determine v,, each test was interrupted at regular increments of strain and the specimen removed from the machine. At the lower strains, when v, was less than about 1 pm, measurements were taken on a Zeiss Linnik interference microscope;

Jan 1, 1969

By C. D. King

DURING the 30-year period spanned by these annual Howe Memorial presentations, many lecturers could proudly claim a kinship either as a student or an associate of the man whose memory we honor. Although it has been my good fortune to have attended many of these annual lectures, it was not my privilege to have known Henry Marion Howe personally. However, his great repute as teacher and scientist was known to all undergraduates of my day and the later years have enhanced my appreciation of his wisdom and foresight. Those who knew him well have said he derived particular pleasure from speculations on the future world of metallurgy. For this reason, I feel that perhaps he would not be unsympathetic to a lecture in his honor which departs from the highly instructive scientific presentations made in the past by so many able Howe Memorial lecturers, and which is concerned more with the practical phases of various steelmak-ing processes and some speculations on their future form and relative importance. The word "revolutionary" is frequently applied to each seemingly important improvement in the production of steel ingots, but in retrospect these changes, impressive as they appear at the time, are merely steps of progress. In the hundred years from the inception of tonnage steelmaking, only three processes can be truly classified as revolutionary. They are the pneumatic process, known in this country as the bessemer process; the reverberatory method called the open hearth process; and, the electric furnace process. There have been many variations and combinations of the three fundamental methods, but they remain truly the only revolutionary methods in steelmaking since its early history. Everything else has been evolutionary, in effect. doing the same things that we have done in the past but doing them better, correcting our errors through experience, and slowly but inevitably reaching a higher state of accomplishment. It has often been said that coming events cast their shadows before, and the production of steel ingots is no exception. As a result of the unrelenting demands of World War I1 and the years that followed, truly impressive progress has been made in steel ingot production. The incessant pressure for immediate results during this period required the employment of initiative and daring, as in few past decades, and many developments were brought to fruition. Of equal importance is the possible effect on future steelmaking methods of the many ideas initiated but still in formative stages. Fig. 1 portrays ingot production in the United States by the three fundamental processes over a period of 75 years and is interesting because it poses some questions as to future trends. The early ascendancy of the bessemer, its replacement in importance by the open hearth process, the amazing growth of the latter, and the recent challenge of the electric furnace are evident from the chart. Management is fully aware of these changes, but is even more interested in the future trends. Our concepts of the relative importance of the more recent developments and their possible effect on future processes may perhaps be best exemplified by a specific, hypothetical problem. Let us assume management is contemplating a new ingot producing plant with an output of 100,000 net tons per month, located in an area where some purchased scrap may be obtained but where by far the largest component will be own-produced blast furnace iron. Management requires a process or combination of processes which will yield highly uniform quality characteristics in the ingot form, and represent the soundest selection in investment and operating cost. Under these conditions, the obvious selection for the past four decades has been the open hearth process but, in view of more recent developments, management may believe that it is no longer permissible to disregard other possibilities with impunity. Accordingly, to be assured of the best possible selection, they request that you review not only the possibilities of utilizing the conventional open hearth, duplex, bessemer, and electric furnace methods, but also the more recent developments, such as the turbo-hearth, the Linz-Donawitz method, the Perrin modifications, and other possibilities. With this background, one might then appraise the relative importance of these methods to meet a specific need, and concurrently speculate on the forms that future ingot processes will assume and the relative importance of these processes.

Jan 1, 1955

By F. B. Sherman, Keros Cartwright

Sanitary land filling has become one of the most widely used methods of disposing of solid refuse. A principal concern of regulatory agencies and the public itself is that landfill operations do not degrade the physical environment, including water resources, and the ground-water reservoir in particular. Knowledge of ground-water and engineering geology can guide landfill operations into suitable terranes or develop measures to compensate for natural limitations at a particular site. Experience and research in Illinois suggest four activities relating to landfill disposal that warrant attention by geologists and engineers: (1) regional delineation of favorable and unfavorable hydrogeologic conditions to facilitate planning and preliminary screening of potential landfill sites; (2) site evaluations, with considerations of geologic materials, topography, water levels, flow systems, and local occurrence and use of water resources; (3) research on aspects of the hydrogeologic environment that control effects of, or are modified by, landfills; and (4) formulation of practices in the siting, construction, and operation of landfills that prevent, mitigate, or isolate deleterious effects. The first two activities are basically in the domain of earth science, requiring the application of fundamental concepts of geology and hydrology and conventional site-exploration methods. The third activity, research, requires contributions from geology as well as other disciplines, including soil physics, sanitary engineering, and chemistry. The fourth calls for policy decisions by regulatory agencies and elected officials, using the contributions of scientists and engineers. Throughout history, man has disposed of unwanted materials by dumping. As urbanization has increased, haphazard dumping practices have given way to disposal under more controlled conditions because of increasing congestion of population and production of waste and greater concern for public health and environmental amenities. Many states and communities have already outlawed open dumping and open burning of refuse. The only practical methods of disposal of large volumes of refuse, therefore, are contained, high-temperature incineration, or burial in a sanitary landfill. In Illinois, regulation of solid waste disposal has been delegated to the Environmental Protection Agency. Each session of the legislature since 1965 has passed increasingly strict laws regulating waste diposal. As a result, the work of evaluating sanitary landfill sites has increased significantly for both the Department of Public Health, now the Environmental Protection Agency, and the Illinois State Geological Survey, which advises the Agency on matters of ground-water geology and pollution. In fact, our ground-water staff spends as much time on studies relating to waste disposal, primarily sanitary landfills, as on ground-water resource studies. Many other geological agencies are experiencing similar demands for increased assistance in solving waste-disposal problems. This paper summarizes some of the salient features of the sanitary landfill concept, describes activities of the Illinois Geological Survey in ground-water and engineering geology relating to landfills, and suggests policies that need consideration. A sanitary landfill is located and operated in such a way that vermin and pests, nuisances, and degradation of air and water are kept at acceptable levels. Some of the physical requirements of a sanitary landfill are all-weather roads for year-round access, fences to retain blowing paper, a daily cover of at least six inches of suitable earth material, and a final cover of at least 2 ft of earth material. Dumping into or adjacent to standing water generally is not allowed. Two common operating techniques are used. In the first, trenches are dug, the refuse is placed in them, and the earth removed from the trenches is used to cover the waste. In the second method, area fill, refuse is placed in low ground and covered with earth from adjacent high areas. The hydrology of the site is a prime consideration in locating sanitary landfills. Putrescible refuse, if saturated above field capacity, produces a leachate that usually has a high concentration of dissolved solids.2 As the leachate also acts as an agent for transporting bacterial pollutants, it constitutes a potential pollution hazard. To reduce the production of leachates, the topography of the landfill area should be such that surface water will not flow into or through the fill. Operations that will result in refuse disposal below or near the highest known water-table elevation may be required to take corrective or preventive measures to protect the ground water. In practice, under humid conditions such as prevail in Illinois, locations where disposal can take place above the water table are relatively few because surficial materials are commonly fine-grained, which permits slow gravity drainage and results in high degrees of saturation (100% moisture content) near the surface. At some sites, although disposal has taken place above the water table, ground-water mounds have developed, resulting in permanent saturation of the refuse. Permeability barriers usually are required to protect the ground-water reservoir from degradation by leachates. The convention in Illinois is to have a minimum of

Jan 1, 1972

By S. Ramachandran, R. G. Woolery

INTRODUCTION Union Carbide is currently operating an in-situ leach project on the Palangana Dome area in Duval county. This deposit meets all the requirements for in-situ leach in that the ore (1) is below the water table, (2) is in a permeable horizon, (3) is amenable to chemical leaching, and (4) is confined by impervious layers. This project has been under commercial production since 1976, and its capacity has been expanded on three occasions since going on-stream. Recently, additional uranium reserves were discovered on the Rogers-Cardenas (R-C) property about 32 km north of the Palangana operation. The ore is located within the Oakville sands and its characteristics are quite similar to those of the ore at Palangana. Both are an unconsolidated Arkosic sand high in clay and calcium carbonate. The R-C ore, however, is somewhat coarser with a mean particle size of 0.15 mm as compared to a mean particle size of 0.07 mm for the Palangana ore. In all respects it would appear that this ore would be a candidate for in-situ leach as a satellite operation to Palangana. Unfortunately, R-C ore is above the water table and, therefore, not amenable to the Palangana practice. Because of the limited known reserves in this deposit, it is readily apparent that conventional mining and milling are out of the question. However, because of its proximity to our Palangana operation, it seemed worthwhile to consider other options. The most viable route based on our past experience was to heap leach the ore. Our recent success at our Gas Hills facility and our Maybell operation, in employing a heap leach practice to our marginal reserves seemed to be a logical approach for processing this ore. Our experiences at both locations are described in "Heap Leaching - A Case History" by R. G. Woolery et al., Mining Engineering, March 1978. In both instances the process is an acid leach circuit and acid consumption averages 20 kg/t H2SO4. A preliminary feasibility study showed that because of the high strip ratio required for the R-C project to be successful, additional ore reserves must be located and that a method of heap leaching with an alkaline circuit would have to be developed. As a result of this paper study, the decision was made to proceed with a program of additional exploration drilling to determine the total ore reserves that could be mined economically. The Mining Department will evaluate each ore zone for cutoff grade, strip ratio, and expected mining cost. At the same time, a laboratory program to evaluate the available core samples for amenability to heap leaching with respect to an estimate of uranium recovery and processing costs was developed. This program is currently in progress, and at this time, we are just completing our process amenability study. BENCH-SCALE EVALUATION OF THE R-C ORE The initial bench-scale slurry leach tests on the R-C ore showed an acid consumption in excess of 200 kg/t H2SO4. These data, of course, discouraged us from considering this process route. Not only would the acid cost be prohibitive, but the gypsum generated by the reaction of the sulfuric acid with the calcium carbonate of the ore would severely effect the percolation of the lixivant. For this reason, the laboratory program was directed toward an alkaline circuit compatible with heap leaching. Because of the proximity of the R-C property to our Palangana operation, it seemed advisable to integrate the processing of this ore into the production at Palangana. Doing so would enable us to bring the R-C property into production by merely enlarging our present facilities at Palangana; otherwise, construction of a grass roots plant would be necessary. Ideally, the simplest method would be to construct the heaps at Palangana and employ an ammonium carbonate/bicarbonate leachant compatible with the in-situ production liquor. The product liquors could then be co-mingled or processed separately as desired. To determine if this goal was practical, samples of the R-C ore were obtained, and a laboratory program initiated. Heap leach amenability testing consisted of preliminary bench-scale evaluation to determine optimum solution strength and ultimate uranium recovery, followed by small column tests to confirm the bench-scale metallurgy and to determine percolation characteristics. These bench-scale tests are being followed by pilot-scale testing approximating field conditions. As expected, the bench-scale tests showed that the dissolution rate is considerably slower for alkaline leach than has been our experience in acid leaching. Because of the slower reaction rates, product liquor grades will be lower than for acid, as greater volumes of solution are required for satisfactory uranium extractions. The greatest influence on reaction times found in the laboratory was the carbonate/bicarbonate strength and oxidant addition. However, the higher salt concentration reduced the efficiency of the IX resin circuit and about 25g/L salt proved to an upper limit compatible with subsequent IX treatment. The oxidant contributed significantly to the early extraction rate but seemed to have only minimal effect on the total practical U308 extracted or the time required to achieve it. This variable will require larger scale testing to determine if the added cost of the oxidant is actually justifiable. Thus, the small-scale laboratory slurry tests, based on the 0.088% U308 sample available, indicate that leaching at 25g/L ammonium carbonate/bicarbonate, with or without oxidant, we might expect an 80-85% U308 extraction on this ore.

Jan 1, 1979

By D. M. Dunbar

LONG before Columbus discovered America the original inhabitants toiled in the copper workings of the Andean Cordillera. Their best diggings appear to have been at Chuquicamata, site of the huge present-day open pit owned by the Chile Exploration Co., subsidiary of Anaconda Copper Mining Co. Located in Antofagasta Province of northern Chile, Chuquicamata is in the Atacama desert, high on the western slope of the Andes at 9500-ft elevation. In the Chuquicamata area the Atacama desert merges into the Cordillera. The climate is extremely arid and precipitation, either in the form of rain or snow, is very infrequent. The temperature varies from 50° to 85°F during the day, and drops to 20° to 40° at night. The Antofagasta and Bolivia Railroad, which connects the Chilean Port of Antofagasta with La Paz, Bolivia, passes within approximately 6 miles of the plant. A branch from the main line to Chuquicamata serves the property for outgoing copper and incoming freight. The distance to Antofagasta by rail is 163 miles. Previous to the Inca conquest, this region was inhabited by the Chuco Indians, a small tribe, descendants of the Aymaras and Quechuas. The Incas, whose capital was located in Cuzco, Peru, gradually extended their empire to the south, reaching at one time to what is now central Chile. It is believed that they built crude primitive furnaces on the banks of the Rio Salado where on a small scale they smelted copper ore from Chuquicamata. At the time of the subsequent Spanish conquest, legend has it that Diego de Almagro, on his return to Cuzco in 1536 after his invasion of Chile, obtained from these furnaces metallic copper for horseshoes. During the period from 1560 to 1879 the high grade veins of the Chuquicamata deposit were worked superficially by the Spaniards and Bolivians. The treaty signed after the war of 1879 recognized the sovereignty of Chile over this territory and Chile, realizing the importance of the district, started to develop it in an orderly manner, although only high grade veins were worked. New Process Is Key Reports on the deposit were brought to the attention of Albert C. Burrage, a Boston lawyer and financier, in 1910. Chuquicamata was described as a huge deposit of ore that was not amenable to the usual treatment by concentrating and smelting since much of the copper occurred as oxides. He had been working on a process for the recovery of copper from sulphide ore which involved a sulphating roast, leaching of the roasted product, and pre-

Jan 1, 1952

By George W. Mitchell

THE property of Eureka Corp. Ltd. is located in the approximate geographic center of Nevada, 2 miles from Eureka, the county seat. The great sources of power, the Colorado, Snake, and Salmon Rivers and the rivers of northern California, are 300 to 500 miles distant, and no lines serve areas closer than 150 miles. Fuel for diesel and steam generation is available in Utah, 300 to 400 miles to the east. Eureka's railhead is 80 miles north where two trunk lines cross the county. A spur line serves Ely, 77 miles east. Good highways connect Eureka to the railheads. Activity in the Eureka mining district began in the early 1870's. The oxidized high grade lead-sil-ver-gold ore terminated against the footwall of the Ruby Hill fault, and in 1890 the main operations ceased. In 1938 Eureka Corp. Ltd. discovered ore in the hanging wall of the fault by diamond drilling. The history of Eureka in the late 1800's indicates that there was some water at 600 to 800 ft in the old workings, probably accumulations above the water table which did not seriously interfere with mining operations. Both the Locan and Richmond shafts were sunk to a level below the table, but apparently the only serious difficulty with water occurred in the Locan. The steam pump used when the last work was done on the 1200 level in 1923, many years after exhaustion of the main orebodies, is still installed on the Locan 900 level. The capacity was about 500 gpm, lifting 750 ft to the 100 level, which connected with the surface. In addition to this. bailers were used to keep the 1200 level free of water. It is said that pumping in 1923 lowered the water in the Holly shaft, about a mile and a half away, but this seems doubtful. The pumping was of short duration because no ore was found. When work at the new Fad shaft was started in 1941 Eureka Corp. Ltd. engineers were fully aware of the probability of encountering water in large volume. Their primary exploration and development had to be carried on at the 2250 level. The first water was encountered at 300 ft. This was undoubtedly surface drainage in the bedding of the Pogonip limestone and was less than 100 gpm. The fractured, loose Hamburg dolomite at the water table was not well cemented, and relatively little water, 300 gpm, percolated through it with difficulty. At 1350 ft well-cemented dolomite containing some open fractures was encountered. These fractures produced the first water of consequence, 750 gpm. At 1700 ft the volume was 1000 gpm increasing to the maximum during shaft sinking, 1600 gpm, at the 2000 level. Secret Canyon shale, a dry formation, was entered at 2100 ft, where a concrete water ring was placed to catch all of the water. The volume decreased rapidly to a constant flow of 1200 gpm. Below 2100 ft the shaft and stations remained in the shale and water was not a problem. Several faults of moderate displacement, including the reverse Martin fault, had been intersected during the traversing of 1000 ft of wet Hamburg, but no undue quantities of water were encountered. Observations in the diamond drill holes in the ore zone area showed a rapid lowering of the water table. The shaft was flooded when it left the dry shale and entered the water-bearing Eldorado dolomite on the 2250 level, crossing a fissure which paralleled the Martin fault. High pressure water doubled the volume then being pumped. Pipe failure through a water door bulkhead was a contributing factor. Immediately following this flooding in March 1948 preparations were made to recover the shaft as rapidly as possible by increasing power and pump capacities as needed. Measurements before flooding indicated the water could be lowered at a fast rate. However, the water table did not recede as rapidly as expected and volumes required to lower the water in the shaft were higher. Obviously the size of the main water channel on the 2250 level was increasing because of erosion, allowing greater volumes to enter the workings and draining beyond the cone originally being drained during shaft sinking. Eroded material was being deposited in the shaft below the 2250 level in serious proportions. In December 1948 a second flooding of the Fad shaft was allowed for the purpose of reassessing existing conditions and studying alternate methods of attack. The detailed geology of the Eureka mining district, see Fig. 1, has been described during the past 75 years by many geologists.' Only the general features and those which seem to affect the drainage problem will be discussed. The old ore zone, mined between 1870 and 1890, is located in a wedge-shaped block of Eldorado dolomite between the footwall of the Ruby Hill fault and the underlying Prospect Mountain quartzite, see Fig. 2. Production of high grade oxidized lead ore containing high values in gold and silver has been variously estimated at $50 to $90 million. The tonnage mined was probably close to 1,500,000, nearly all of which was found above the water table. The new ore, discovered by diamond drilling in the hanging wall of the Ruby Hill fault, is a flat-

Jan 1, 1954

By R. A. Burton, J. T. Hansford, D. A. T. Donohue

Substantial evidence indicates that many petroleum producing horizons contain naturally occurring, ordered fracture systems and that within a particular geologic zone, vertical fractures induced in wellbores often will be directed along a particular compass direction. Both conditions will seriously alter the fluid displacement behavior within reservoirs. In this study the effect of induced fracture orientation and length on sweep efficiency is determined for a five-spot pattern. In general, it is assumed that all wells are fractured and directed along the same compass direction. Using the electrical analog to steady state, two-dimensional fluid flow in porous media, boundary conditions are obtained from which flood fronts are tracked numerically. The numerical computations require a particle tracking routine for approximating flood front histories. It is shown that recovery is sensitive to the length and orientation of fractures for the pattern studied. with the proper design of fracture-pattern systems, recovery can be enhanced considerably. INTRODUCTION Hydraulic fracturing introduced in 1949l gave the industry a rather inexpensive means of increasing the fluid injection or production capacity of wells. It has been used with particular success to increase the production rate of wells completed in tight formations, such as in western pennsylvania2 where producers have fractured in depleted or near-depleted fields and observed economic responses. Once the natural energy declines in such a reservoir where all wells have been fractured, waterflooding is generally suggested as means of further increasing recovery. Of the dual objective sought in waterflooding — high injectivity and high breakthrough sweep efficiency — the former condition can be obtained if all wells in the flood pattern are fractured; the latter condition should depend on the nature of the fracture system. Considerable theoretical work has been published on the nature of fractures induced in boreholes.3-5 Although discussion persists concerning the possibility of forming a horizontal fracture at a given point within the wellbore, it is generally conceded that only vertical fractures will develop below a given depth, i.e., where the fracturing Pressure is less than the overburden load. Given the fact that fractures will be vertical in most cases of interest, it is also important to know whether there is order to fracture orientations within a given geological region. Kehle has suggested that in tectonically relaxed areas of uncomplicated geology, the stresses are fairly uniform and all fractures in the region should be parallel.5 Dunlap arrived at a similar conclusion in a theoretical investigation of localized stress conditions surrounding the borehole.6 He concluded that most vertical fractures are propagated in a preferred azimuthal direction. Fraser and Pettitt, in extending these theoretical suggestions to a specific field case, used an impression packer to record both a vertical fracture and the orientation of this fracture in the wellbore of a well in the Howard Glasscock field, Tex.7 Use of this information enhanced the waterflood recovery of the field. Anderson and Stahl also used impression packers on three fractured wells in the Allegheny field. N. Y.. and found that the fractures were oriented more or less along the same compass direction.8 Orientation of the fractures in this manner depends on the stress condition within the formation during fracturing. Elkins and Skov have demonstrated that a natural, oriented, vertical fracture system exists within the Spraberry field? Heck suggests

Jan 1, 1969

By Gary E. Butts, Paul R. Smith

Introduction The design of a mineral processing flowsheet involves many factors. The importance of the sample can not be overemphasized. Competent and thorough mineralogical examinations are important as an aid in designing the test program. Designing and conducting the laboratory test program is basic to the design of the flowsheets. In the project under consideration a flotation pilot plant was conducted to prove the process and to provide a realistic prediction of metallurgical results. Moreover, the pilot plant operation provides important pilot plant design data. Finally, accurate and clear reporting is essential so that the information developed can be utilized advantageously by the client and the engineering firm. The design of a minerals processing flowsheet is a team effort. Engineers from the mining company, the research organization, and the engineering firm all contribute to the effort.

Jan 1, 1978

By Forbes K. Wilson

"For his vision, determination, dedication, and leadership in the conquest of the remote and rugged Ertsberg and the technical and human barriers to its development and for his lecture ..." It is a great honor for me to be included among the Jackling Award recipients, many of whom have been close friends and associates. My subject is the development of the Ertsberg, a copper mine located 4' south of the equator, 70 miles (113 km) from the south coast of Irian Jaya---the easternmost province of the republic of Indonesia at an elevation of 12,000 ft (3658 m) in an area which was inhabited by a few "stone age" people when I first arrived there in 1960. Persistence, tenacity, immagination and innovation were all required to bring the Ertsberg into production. The technical, financial and political problems that have been faced over the past 16 years were unique in many respects. I hope that my paper will give some encouragement and help to future developers of mines in difficult and inhospitable surroundings.

Jan 1, 1978

By B. Chalmers

THE practical importance of the phenomena of melting and freezing must have been recognized for a very long time. The difference between ice and water, for example, has had a profound influence on the history of mankind and the evolution of society. The possibility of melting a metal and allowing it to freeze in a mold of chosen shape has been an essential ingredient in our mastery of the art of shaping metals, and therefore in the evolution of the: machine age in which we find ourselves. The importance of melting and freezing, as applied to metals and alloys, has been so great, in fact, that empirical solutions have been found for the multitude of practical problems that have arisen. This approach has been so successful that relatively little attention has been directed to arriving at an understanding of the fundamentals of the processes. But metallurgy has come to a stage at which we may expect that some, at least, of the more complex problems that have not yet been solved (or perhaps even recognized) may be handled more effectively by scientific study, theoretical understanding, and logical experimentation than by trial and error. In this lecture, therefore, I propose to describe in outline what I think really happens when a metal freezes. In doing so I hope to explain many of the phenomena which have been observed, and in particular to account for the structures that are obtained in actual ingots and castings. The basic problem, to which this lecture represents a tentative partial answer, is this: a mass of metal, containing known proportions of various elements, is melted, heated to a given temperature, and then allowed to freeze under specified conditions. What will be the "structure" of the resulting metal? The term structure includes: 1—crystal size, shape, and orientations, 2—distribution of chemical elements, and 3-—shape, including cracks, cavities, pores, etc. The Solid-Liquid Interface We will first consider what takes place if a single crystal of a metal in the form of a rod is heated, not uniformly, but so that one end is hotter than the other. If this heating process is continued long enough, the hotter end will eventually melt; we will suppose that the rod is in a containing vessel so that the molten metal does not run away, Fig. 1. When some of the metal has melted, we have some solid, some liquid, and an interface or surface of contact between them. If the source of heat is now removed, the interface will move so that some of the liquid freezes, and if the supply of heat is suitably adjusted the interface will remain at rest. This very simple arrangement allows us to study the basic processes of melting and freezing, and if we fully understand this simple case, we may be able to account for what takes place under practical conditions where the heat does not all flow in the same direction, and where the heat flow is determined not by a controllable source of heat but by the heat capacity and temperature of metal and mold, and by the heat loss from the mold surface. The solid-liquid interface is evidently the region of the greatest interest to us; on one side of it there is crystalline solid, and on the other, liquid. In the solid, each atom has a well defined position, around which it vibrates as a result of thermal agitation. It only leaves this position in the relatively rare event of a "diffusion jump." The liquid is much less systematically organized. The atoms are about as far from their neighbors as in the solid, but the arrangement is much less systematic and is continuously changing. The solid and the liquid are represented diagrammatically in Fig. 2. The average energy of the atoms in the liquid is greater than in the solid by an amount that corresponds to the latent heat of fusion, i.e., the amount of heat that has to be supplied to convert unit mass of solid into liquid at the same temperature. The Two Processes As has recently been shown by Jackson and Chalmers,3 many of the features of the processes of freezing and melting can be understood if it is assumed that a continuous and rapid interchange of atoms between solid and liquid always takes place at a solid-liquid interface." It is necessary to con- sider two distinct processes, that of melting, in which atoms leave the surface of the solid and become part of the liquid, and the converse process,

Jan 1, 1955

By R. B. Fuller

DURING the summer of 1949, the United Nations Scientific Conference on the Conservation and Utilization of Resources met at Lake Success. As summed up by one writer, the purpose was: "That everyone will try to learn how to feed more people from an acre of land." That objective is a stimulus to the phosphate rock industry as well as other industries.

Jan 8, 1951

By John S. Brown

IN 1925 the writer undertook a detailed statistical study of all producing areas in the Joplin district as a basis for evaluating programs and measuring objectives. For this purpose, the published figures in the yearly volumes of Mineral Resources were used, supplemented for earlier years by publications of the Missouri Geological Survey and other local and less official sources. When all else failed, the available data were projected backward to hazard a reasonable guess as to the unrecorded early output of important areas. Fortunately, the proportion of such prehistory production is not a large factor in any of the totals. These results were used during the next few years to measure the relative importance of various producing areas and to predict the peak period of development of the all-important Picher field. For the purpose of this review, the charts have been completed to the end of 1950. During World War 11, the U. S. Bureau of Mines became interested in a similar study and issued comprehensive statistical tabulations of data up to 1945 ( Info. Circular 7383), which have been checked against the figures used herein. This tabulation, however, does not include all the earlier data used by the writer nor does it offer any estimates of the wholly unrecorded era in the beginnings of the earlier camps.

Jan 9, 1951

By R. V. Norris

1. INTRODUCTION. THE anthracite coal trade, with a shipment averaging about 70,000,000 tons per year, differs essentially from other coal business, in the fact that the larger sizes, comprising about 65 per cent. of the total, are used almost exclusively for domestic purposes, principally during the winter months; and that for proper combustion close-sizing is imperative, so that eight sizes are made; broken, egg, stove, and nut, known as prepared sizes, and pea, buckwheat, rice, and barley, known as steam sizes. These sizes are made in the regular course of preparation, and but little variation in the natural percentage of each is practicable. Thus it is necessary either to work the collieries intermittently, or to dispose of all the varying sizes in their fixed proportions. Unfortunately, the market does not at all times absorb the coal in proper proportions, and rather than interfere with the regular operation of the mines it has been found economical to provide storage for the excess. While the cellars of the consumers are the great storage-plants of the country, this capacity is not under the control of the trade, except so far as reduced prices during the spring and summer months may tempt the use of this reserve. The same remark applies but with less force to the retail yards, which,, though usually small, store in the aggregate a large amount of anthracite. While every effort is made to utilize to the utmost the individual storage-capacity of the country, there still remains the necessity for taking care of the irregularities of the market by the construction of storage-plants under the direct control of the producers. Such plants are usually the property of the railroads, and are situated at points most convenient from a traffic stand-point.

Jun 1, 1911

By AIME AIME

FRANK T. SISCO is a native of Lawrence, Kansas. He received his earlier education in Clinton, Iowa and his university training at the University of Illinois. His experience in the steel industry has included service in the chemical, testing, and metallurgical departments of the Illinois Steel Co., of South Chicago; the Hess Steel Corp., of Baltimore; the IJ. S. Naval Ordnance Plant at Charleston, W. Va.; and the American Steel and Wire Co., Waukegan, Ill. He severed for seven years as chief of the metallurgical laboratories of the U. S. Army Air Corps at Wright Field, Dayton, before coming to New York in 1930 to become metallurgist and editor of the Alloys of Iron Research, the position he now holds.

Jan 1, 1940

By J. J. Reed

The mines of southeast Missouri's Lead Belt have been in operation since 1864, almost 100 years. During this period about 10 pct of the total ore available has been left in place as pillars, and this now represents the best ore reserve left to the older mines. St. Joseph Lead Co. has been investigating various plans of pillar recovery for several years. The example to be described is only one possible approach under a specific set of conditions. Conditions vary greatly in the different areas within the Lead Belt, and the same approach would be entirely inadequate in certain other areas, but there are important sections where this system is clearly applicable. Horizon mining is practiced in the Lead Belt, and stopes may vary from 10 ft to well over 100 ft high. Fig. 1 is a plan view of 6100 stope, the area to be discussed. For these mines this is a small stope, only 10 to 12 ft high. Before any plans were drawn up for pillar recovery, three independent estimates of the pillar grades were made by the mine superintendent, the mine captain, and the mine geologist. After these results were averaged, pillars indicated to be of economic grade were marked for removal as shown. At this stage several alternatives were considered: 1) retreat with caving, 2) concrete replacement pillars, and 3) systematic bolting and retreat. A retreating system with caving was not chosen because the roof was known to be too bad even to permit loading a blasted pillar before caving would begin. Concrete replacement pillars would have been too costly and would have required a large crew, bulky materials, and unwieldy equipment. Systematic bolting was chosen as being most acceptable for the following reasons: 1) Limited area and spans were involved. 2) Only minor modifications to the jumbo would be required for bolting. 3) Only a two-man bolting crew would be needed. 4) Only compact bolting materials would be required. 5) The bolting plan could be modified to allow

Jan 1, 1960

THE compiler, or author, of the Probierbücblein is not known, and the date of its first printing is uncertain. It was obviously not written by an assayer who wished to give a coherent description of his craft for the instruction of others; this was left for Biringuccio first to do. It is easy to believe that originally the Probierbücblein was a random collection from various sources of information useful to an assayer, probably gathered by one of the profession for his own use. It must remain a matter of speculation whether this was shown to a printer, who recognized the opportunity to turn an honest penny and made an inept attempt to give the manuscript the appearance of a coherent account by writing a preface and arranging the various items in somewhat logical sequence, or whether the assayer him- self started this job and was prevented from finishing it. That some sections of the book, especially near the be ginning, present an almost continuous story, while others are chaotic in their arrangement; that several paragraphs are printed more than once in different places; that there are paragraph headings, for which the text was never written, or lost, followed by irrelevant information-all this seems evidence of a job unfinished and points to a publisher's offering for the public's quick money. The demand must have been brisk, for metal working was a profitable profession, and the jeweler who was working with metals but did not care to become completely trained in assay methods would jump at the chance of acquiring easy knowledge. Even the great Agricola has whole sections obviously plagiarized from the Probierbücblein. Neither Biringuccio nor Ercker shows any direct influence, though they and all subsequent metallurgists will in some degree be affected by

Jan 1, 1949

By R. R. Estill

THE rank of the Ruhr coal ranges from a high volatile bituminous coal to an anthracite, depending to some extent on the original depth of the seam. The average Ruhr coal corresponds to a soft bituminous American coal of a coking quality. The average thicknesses of individual coal seams being mined are also comparable (59 in. against 65 in. in the United States). However, consideration of seam conditions and mining conditions other than those just mentioned emphasizes differences rather than similarities with United States soft coal. In general, the Ruhr seams now being mined are much more folded and inclined than American seams. Dips of 20' and 30" are common in seams now being worked, and 30 pct of the coal reserves in the district are in seams dipping more than 35". Only on the tops and bottoms of folds do we find rather flat coal seams. In addition to the folding there is extensive displacement by cross faulting plus a certain amount of strike faulting of an overthrust nature, which results locally in doubling or omission of seams. Because of the long history of mining in the Ruhr, nearly all coal lying near the surface has long since been mined out, and we find that the average depth of mining is at present about 2300 ft below the surface. Deep mining, folding, and faulting result in seam conditions requiring a great deal more roof support than one finds in American soft coal mines. In fact only in the anthracite district and the Rocky Mountain and Pacific coal fields do we find somewhat similar conditions. It is easy to say, therefore, that the problem of mechanization of coal cutting and loading in the German mines is quite different from that which we have so effectively met in America with our mobile cutters and loaders, duck bill loaders, and a room and pillar system of mining our drift and slope mines. Partly because of more limited coal reserves, the traditional German mining system is largely the longwall method, which gives an almost complete coal recovery. Backfilling must be extensively practiced to protect the longwall faces, the over and underlying seams and workings, and especially the surface industrialized areas and barge canals. The German engineers have accordingly concentrated their efforts on the design of cutters, loaders, and conveyors suitable to longwall methods rather than room and pillar methods. Undercutters with cutter bars like American models have been in use in the Ruhr since well before World War 11. In 1941 they accounted for 8.5 pct of the production. This percentage, of course, includes coal which was undercut but nevertheless had to be broken down with air hammers or with explosives. The most common of these cutters is the Eickhoff Standard cutter (see fig. 1). This machine does about 95 pct of the undercutting in the Ruhr today, and is available with either compressed air or electrical power and in at least four different sizes. A variation of the cutter is this one with two cutter bars (fig. 2). At the end of 1947 about 200 of these machines and similar cutters were accounting for 13.2 pct of the total production, a production which was, however, only 60 pct of the 1941 production rate, so that the actual cutter tonnage was only up to a small amount over 1941. In 1941 about 3 pct of the production was accounted for by shearing machines making their cut perpendicular to the longwall face. They were similar to those used in the States. These machines are today considered obsolete and now account for only 0.7 pct of the total production. They are located at only a few mines and at present do not seem to have much of a future in the Ruhr. For the future, the Ruhr miner is looking forward to rather extensive mechanization of face work, with two major types of equipment being developed almost simultaneously. On one hand there is the development of cutter loaders for use in relatively hard coal. They represent the further extension of ideas developed after relatively long experience with the Eickhoff cutter. On the other hand there has been since 1942 an intense interest in the Ruhr in the development of face-stripping methods, particularly by the Kohlenhobel (coal plow) and its modification. At the end of 1947 these cutter loaders, Kohlen-hobels and scrapers together were actually accounting for only about 1.4 pct of total production while air hammers still broke 77.1 pct and as much as 1.2 pct was actually broken by hand picks. However,

Jan 1, 1951

By R. R. Estill

THE rank of the Ruhr coal ranges from a high volatile bituminous coal to an anthracite, depending to some extent on the original depth of the seam. The average Ruhr coal corresponds to a soft bituminous American coal of a coking quality. The average thicknesses of individual coal seams being mined are also comparable (59 in. against 65 in. in the United States). However, consideration of seam conditions and mining conditions other than those just mentioned emphasizes differences rather than similarities with United States soft coal. In general, the Ruhr seams now being mined are much more folded and inclined than American seams. Dips of 20' and 30" are common in seams now being worked, and 30 pct of the coal reserves in the district are in seams dipping more than 35". Only on the tops and bottoms of folds do we find rather flat coal seams. In addition to the folding there is extensive displacement by cross faulting plus a certain amount of strike faulting of an overthrust nature, which results locally in doubling or omission of seams. Because of the long history of mining in the Ruhr, nearly all coal lying near the surface has long since been mined out, and we find that the average depth of mining is at present about 2300 ft below the surface. Deep mining, folding, and faulting result in seam conditions requiring a great deal more roof support than one finds in American soft coal mines. In fact only in the anthracite district and the Rocky Mountain and Pacific coal fields do we find somewhat similar conditions. It is easy to say, therefore, that the problem of mechanization of coal cutting and loading in the German mines is quite different from that which we have so effectively met in America with our mobile cutters and loaders, duck bill loaders, and a room and pillar system of mining our drift and slope mines. Partly because of more limited coal reserves, the traditional German mining system is largely the longwall method, which gives an almost complete coal recovery. Backfilling must be extensively practiced to protect the longwall faces, the over and underlying seams and workings, and especially the surface industrialized areas and barge canals. The German engineers have accordingly concentrated their efforts on the design of cutters, loaders, and conveyors suitable to longwall methods rather than room and pillar methods. Undercutters with cutter bars like American models have been in use in the Ruhr since well before World War 11. In 1941 they accounted for 8.5 pct of the production. This percentage, of course, includes coal which was undercut but nevertheless had to be broken down with air hammers or with explosives. The most common of these cutters is the Eickhoff Standard cutter (see fig. 1). This machine does about 95 pct of the undercutting in the Ruhr today, and is available with either compressed air or electrical power and in at least four different sizes. A variation of the cutter is this one with two cutter bars (fig. 2). At the end of 1947 about 200 of these machines and similar cutters were accounting for 13.2 pct of the total production, a production which was, however, only 60 pct of the 1941 production rate, so that the actual cutter tonnage was only up to a small amount over 1941. In 1941 about 3 pct of the production was accounted for by shearing machines making their cut perpendicular to the longwall face. They were similar to those used in the States. These machines are today considered obsolete and now account for only 0.7 pct of the total production. They are located at only a few mines and at present do not seem to have much of a future in the Ruhr. For the future, the Ruhr miner is looking forward to rather extensive mechanization of face work, with two major types of equipment being developed almost simultaneously. On one hand there is the development of cutter loaders for use in relatively hard coal. They represent the further extension of ideas developed after relatively long experience with the Eickhoff cutter. On the other hand there has been since 1942 an intense interest in the Ruhr in the development of face-stripping methods, particularly by the Kohlenhobel (coal plow) and its modification. At the end of 1947 these cutter loaders, Kohlen-hobels and scrapers together were actually accounting for only about 1.4 pct of total production while air hammers still broke 77.1 pct and as much as 1.2 pct was actually broken by hand picks. However,

Jan 1, 1951

By R. A. Vandermeer, J. C. Ogle

The preferred crystallographic orientations (texture) developed in randomly oriented, poly crystalline niobium during rolling were studied by means of X-ray diflraction techniques. The evolution of texture at both the surface and center regions of the rolled strip was carefully examined as a function of increasing defamation throughout the range 43 to 99.5 pct reduction in thickness. Certain aspects of the center texture development in niobium are in agreement with the predictions of a theory by Dillamore and Roberts, but others cannot be explained by the theory in its present form. Above 87 pct reduction by rolling, a distinctly different texture appeared in the surface layers which was unlike the center texture. The present results are compared with previous results obtained from other bcc metals and alloys. RANDOMLY oriented, poly crystalline metal aggregates when plastically deformed to a sufficiently large extent develop preferred orientations or textures. In a recent review article, Dillamore and Roberts1 pointed out that the nature of the developed texture may be influenced by a large number of variables. These include both material variables such as crystal structure and composition and treatment variables such as stress system, amount of deformation, deformation temperature, strain rate, prior thermal-mechanical history, and so forth. From a practical point of view, the control of preferred orientation may often be important for the successful fabrication of metals into usable components. During the past few decades many experiments have been devoted to the study of deformation textures. This work, however, has been confined in large part to metals and alloys that have an fcc crystal lattice. By comparison, bcc metals and alloys have received much less attention, and consequently our understanding of preferred orientations in these materials is only shallow. This state of affairs worsens when it is realized that almost all of our present howledge about this class of materials derives from studies on irons and steels.' The bcc refractory metals, which are relative newcomers to the industrial world, have, on the other hand, been given at best only passing glances in the area of texture development. Our understanding of the evolution of preferred orientations in bcc metals can only remain fairly limited until systematic studies of metals and alloys other than the irons and steels have been carried out and the influence of the many variables has been determined. To that end a program was initiated to investigate in detail texture development in niobium. The present paper reports some of the results of this study. Textures were determined at both the center and surface of strips rolled variously to as much as 99.5 pct reduction in thickness at subzero temperatures. Emphasis in this paper is on texture description and on texture evolution during rolling to progressively heavier deformation. EXPERIMENTAL PROCEDURE The niobium was purchased from the Wah Chang Corp. as a 3-in.-diam electron-beam-melted billet. Chemical analysis indicated the impurities to be less than 300 ppm Ta, 40 ppm C, 10 ppm H, 170 ppm 0, and 110 ppm N. All other impurities were below the limits of detection by spectrochemical analysis. This large-grained billet was fabricated into specimen stock so that a fine-grained randomly oriented grain structure resulted. This was accomplished in three deformation steps alternated with recrystalli-zation anneals of 1 hr at 1200°C in a vacuum of low 10"6 Torr range after each deformation step. The first step was to alternately compress the billet 10 to 20 pct in each of three orthogonal directions. The second step was to compress in only two directions 90 deg apart to produce a 2-in.-sq bar. The final step was to roll this bar 50 pct to give a 1-in. by 2-in. cross section. After the final anneal, metallo-graphic examination showed the material to have an average grain size equivalent to ASTM No. 5 at 100 times (i.e., 0.065 in. diam). Specimens cut from the center and edges of this bar gave no indication of detectable preferred orientation when examined by X-ray diffraction. Samples 1.5 in. long, either 0.625 or 0.750 in. wide, and approximately 0.400 in. thick were machined from this fabricated ingot. The surfaces corresponding to the rolling planes were ground so as to be parallel. The samples were chemically polished in a solution of 60 pct nitric acid and 40 pct hydrofluoric acid (48 pct solution) prior to rolling to remove any cold work introduced in the machining operations. Rolling was accomplished with a 2-high hand-operated laboratory rolling mill that had 2.72-in.-diam rolls. Prior to operation, the rolls were polished with 600 grit paper, cleaned with acetone, and then soaked in a container of liquid nitrogen for several hours. The samples were also soaked in liquid nitrogen prior to rolling and were recooled between each pass. While some slight heating of the samples occurred during rolling, this procedure maintained the sample temperature well below 0°C at all times. The samples were rolled unidirectionally, and the rolling plane surfaces were not inverted during any phase of the operation. The draft per pass averaged between 0.010 to 0.012 in. After 96 or 97 pct reduction the draft was reduced to 0.001 to 0.002 in. per pass. Samples were rolled to various reductions in thickness between 43 and 99.5 pct.

Jan 1, 1969

By Peter J. Conroy, Julianne H. Gyarmaty

INTRODUCTION The current study is part of the U.S. Department of Energy's (DOE) on-going subsidence research program. The long-term objective of the DOE program is to develop analytical methods of subsidence prediction applicable over a broad range of geologic and mining conditions encountered in the United States. The DOE currently has several contracts in progress to develop a sound data base that can be used in the development of predictive models. Dames & Moore is currently under contract to DOE to collect and analyze detailed subsidence and hydrologic data from a coal mine employing longwall mining methods in the Eastern Coal Province. Work under Contract No. DE-AC22-80PC 30335 began in October of 1980 and will continue through September of 1984. The contract includes monitoring of two adjacent longwall panels. This paper summarizes the surface monitoring work that was performed through February of 1982. Mining of the first of the two panels (3 off 3 north) was completed on February 3, 1982. Results of the surface monitoring through completion of this panel are presented and discussed. SITE DESCRIPTION The mine being studied is located in eastern Ohio. The site is west of the Ohio River, approximately 56 km (35 miles) south of Wheeling, West Virginia. The topography of the unglaciated region

Jan 1, 1982