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By G. A. Vissac

O. R. LYONS *—I wish to thank Mr. Vissac for his compliment. I hope that his paper is not only well received, but that it will serve to bring forth more papers on the subject of thermal drying. One of the primary purposes of the work performed by Battelle for Bituminous Coal Research in investigating the thermal drying of coal was to stimulate other investigators and to get them to contribute their knowledge in the form of papers such as this one. We at Battelle and the personnel of Bituminous Coal Research are very gratified that Mr. Vissac and other persons have responded in this matter of the thermal drying of coal. I wish to state that I think that Mr. Vissac's paper is a very clear and easily understood description of a method of calculating the design requirements for a screen type drier, and I think that it would be exceedingly valuable to operators and to those who intend to purchase any type of thermal drier and use it in the future, if the manufacturers or operators who have such information for other types of driers would provide the same type of information for the other makes of driers now on the market. 1 also wish to point out—an idea that is new to me, and I know is new to most of the operators of driers in the United States-—the idea of recovering the heat that is normally lost in the coal and in the exhaust gases. This heat is not being recovered at most (of the thermal drying operations in the United States, and the possibility of recovering it should be called to the attention of every single one of those operators. I know many of them have never given any thought to the matter, but they will be interested once they realize the ease with which it could be done and the savings that could be realized. I also wish to compliment Mr. Vissac for presenting the method of analysis that he uses to determine the difficulty of drying any particular coal. It is a very simple method, and yet it seems to me that it should be a very effective, very efficient method for determining the difficulty of drying for his particular problems. C. Y. HEINER*—I do not know that I can add anything very illuminating to what Mr. Vissac has said. I think anything that Mr. Vissac said in regard to coal drying is a contribution because, to my personal knowledge, he has studied the matter carefully for many years and made many valuable contributions. I am not too familiar with coal drying problems in the east, but I know in the west we have not made enough coal drying studies. I think coal operators too often just take the coal as it is and make more or less the best of it. There are relatively few washing plants in the west now, and so the problem has not come to the front as much as it probably will in the future. In this connection, it seems to me that this matter of drying the raw coal, as Mr. Vissac brings up, is an extremely important one. We have not a continuous miner ourselves, yet, but we expect to get some this year, and we think the percentage of fine coal-—that is, minus 3/16 in.—will double. We have about 20 pct minus 3/16 in. in the 8 in. by 0 size now, and we think we will likely have 40 pct, which will have a surface moisture of the order of 8 pct. To wash it satisfactorily, we will have to dry the raw coal first in order to screen it, and after that, I suppose, there will have to be dry cleaning of some sort. We have not really used dry cleaning on fines in the west yet to my knowledge, but it is a matter that has to be faced by the industry, and I am very hopeful that Mr. Vissac's study will assist us in that connection. W. L. McMORRIS*-In my company we are preparing largely metallurgical coal for a great number of byproduct coke plants. The most outstanding thing to me about the requirements of moisture in the finished product is that there is a different requirement for almost every coke plant. Each operator has a different set of factors on which he establishes his coking costs where they involve moisture. For our corporation operations in Birmingham, my company does not produce the coal, but in Birmingham they are getting away with moistures very much higher than our plant at Clairton, Pa., would tolerate. The moisture that we have to produce for the plants along the lakefront where they are subject to much more severe weather is something else again. We have not tackled heat drying, primarily because our customers do not know what heat drying will do to the coking characteristics of the coal. If the temperature of drying can be held down

Jan 1, 1950

By H. E. Stauss, J. Hino

INTEREST in silicon has arisen again in the past decade as a result of improvements in crystal rectifiers.' Although the preparation of silicon was first reported by Berzelius in 1880, the early product was of relatively low purity, and only the need for rectifiers in World War II led to the production of a 99.9+ pct pure powder. This material in crystalline form was consolidated into massive silicon for use, and the method developed was to melt it with selected added constituents as "doping" agents. Melting techniques, therefore, are of great importance. There are two basic problems in producing silicon ingots free of doping additions; one is the prevention of spitting and the other is prevention of cracking of the ingot during freezing. The most satisfactory arrangement yet developed for producing massive silicon is to melt and freeze in a cylindrical quartz crucible surrounded by a concentric heating element and concentric radiation shields or insulation. For example, use can be made of a tubular heater with a high frequency generator as the source of power and reflecting shields of alundum cylinders. The spitting of silicon is related to gas evolution, and the gas comes from two primary causes—adsorbed gas and the reaction products of silicon and the crucible. Gas is also released from bubbles contained in the quartz crucible walls. Improved removal of adsorbed gas can be achieved by means of controlled melting and freezing. The seriousness of the problem in vacuo is reduced with an electrically operated mechanical movement of the high frequency power coil. The upper portion of the powder charge is melted first and the high frequency coil lowered until the powder is completely molten. During cooling the high frequency coil is raised slowly. These means also reduce the final nonviolent extrusion of large beads of metal through the ingot top during freezing. Better control of spitting and bead extrusion is obtained when melting is done under helium at. atmospheric pressure instead of in vacuo. The problem of reaction between silicon charge and crucible in practice is confined to the reaction between silicon and quartz. This2 apparently is: Si + SiO2 + 2SiO The part that this reaction plays in spitting has not been isolated for separate study. SiO is a volatile vapor at the melting point; of silicon and is released freely during melting in vacuo, but hardly at all in helium at atmospheric pressure. The cracking of ingots is a major difficulty in melting silicon, and its prevention requires special melting techniques or the addition of "toughening" agents such as aluminum or beryllium.' The cracking of the ingots has been explained as being the result of the expansion that occurs upon freezing; although direct observation of freezing ingots reveals visible cracks on the surface only after a red heat has been reached, suggesting that cracking is the result of differential contraction of silicon and quartz. Silicon wets quartz, and the ingot adheres tightly to the crucible. Therefore as ingot and crucible cool, the two either have to pull apart, or at least one must crack. Surprisingly, in spite of the relative thinness of the quartz and the thickness of the ingot, the ingot and the crucible both crack. Microscopic and X-ray4 studies fail to show any plastic flow other than twinning in the ingots. Slow cooling fails to prevent cracking. Another possible solution to cracking is to weaken the crucible. Use of thin-walled crucibles finally led to success with fused quartz crucibles with a wall thickness of 0.25 to 0.50 mm. With such thin-walled fused quartz crucibles consistently uniform success is secured in producing sound ingots 30 mm in diam from the purest available grade of silicon (99.9+) without the use of any type of addition. Melts are made in the size range of 50 to 100 g. Omission of a deliberately added doping agent is not sufficient to insure pure ingots. The reaction of silicon with crucibles and the resultant solution of impurities in the silicon is well-established." In this laboratory, the presence of Al, Be, and Zr has been found spectroscopically in ingots melted in contact with alumina, beryllia, and zircon. The best crucible materials reported in the literature are MgO and SiO2. Use of MgO in this laboratory has resulted in a heavy deposit of magnesium on the furnace walls, showing that a reduction of the magnesia occurred and the resulting magnesium removed from the melt by volatilization. In the case of quartz, the silica is reduced and SiO liberated to deposit on the equipment walls. There probably is real danger that oxygen is dissolved in the ingot when either magnesia or silica is used as the crucible material. Preliminary analyses by Dean Walter in his vacuum unit in this laboratory6 indicate the presence of oxygen in undoped silicon melted in quartz.

Jan 1, 1953

By J. C. Swartz

New measurements of graphite and cementite solubilities in ferrite have been obtained to resolve disagreements among previous data. To measure graphite solubility, specimens of iron were equilibrated with rarious carburizing atmospheres at temperatures be-twee~z 480" and 1435°C. The carbon activity of each atmosphere relative to graphite was taken from published data. After quenching from each equilibration treatment the concentration of carbon (typically 10 pprn by wt) was determined from the height of the Snoek peak in the internal friction spectrum. A few inclividual analyses by combustion support the reszdts of Smith on the solubility of graphite near the eutectoid temperature. Combination of these combustion and internal friction analyses determines In a, (at 0.01 wt pct) = 24,000/RT - 12.72, where a, is the activtity relative to graphite in ferrite at fixed cornposition. The same specimens were used to determine the solubility of self-stressed cementite precipitates. The results indicate solubilities of 6.3, 23.2, and 68 pprn carbon by wt at respectively 400°, 500". and 600°C. The once large discrepancy between the cementite solubility determined by internal frictiori and that calculated from flee energy data has been almost completely resolved. A recent study1 (hereafter designated I) on the solubility of cementite in a iron did not fully resolve the disagreement among previous data. The remaining disagreement is illustrated by the phase diagram in Fig. L The data of I are represented by the line labeled "stress-free" cementite. They determine a heat of solution of cementite of (14.8 * 0.2) kcal per mole which contrasts with the (20 * 1) kcal per mole derived from Smith's graphite solubility2 and the formation energy of cementite.3 Toward resolving this disagreement we have obtained new data on the solubilities of graphite and cementite. METHODS As in I the Snoek peak in the internal friction spectrum was used to measure the concentration of carbon dissolved in ferrite. This method relies on the relation C=kQ-1 [1] in which C is the concentration of dissolved carbon, Q-I is the Snoek peak height, and k is a proportionality factor which is a function of the peak temperature, polycrystalline texture, and other parameters characteristic of the ferrite lattice. To determine true heats of solution via Eq. [1], it is necessary to establish that k is independent of the temperature from which the specimen is quenched. Some evidence on this point is contained in a separate report.* Additional evidence is shown by the data points in Fig. 1. These represent the Snoek peaks of two specimens carburized to about 10 ppm C, copper plated, and quenched first from 888º then from 588°C. The copper plate was removed before each Snoek peak measurement. These data show that there is no detectable dependence of k on quenching temperature. The solubility of graphite was determined by repeatedly carburizing various specimens of iron wire with gas mixtures of known carbon activity relative to graphite. To cover a wide temperature range within the a phase boundaries it is desirable to use as low a carbon concentration as is compatible with reasonable precision. With the present internal friction apparatus the best compromise of these desires is a concentration near 10 wt ppm. The dashed line in Fig. 1 was used to approximate the activity of carbon* for

Jan 1, 1970

By A. J. Lynch, M. J. Lees

Introduction The design of wet grinding circuits may vary considerably according to the duty which is to be performed. Grinding machines which are commonly used in these circuits are rod mills, ball mills, autoge¬nous mills, and pebble mills while size separation machines commonly used are vibrating and wedge-wire screens, hydrocyclones, and rake classifiers. In some circuits, such as those producing feed to cement kilns, all the ore entering the circuit is discharged in a single product stream. In others, such as those operating on magnetite-taconite ores, the feed entering the circuit leaves in two or more streams. Grinding circuits in flotation plants normally include sections in which either the concentrate or the tailings from the primary flotation circuit is reground and these additional sections should be regarded as an inte¬gral part of the grinding circuit. A typical circuit is shown in Fig. 18. Problems in achieving maximum efficiency may occur at three stages: 1) At the design stage when the optimum size reduction flow¬sheet must be selected with respect to the types, numbers, and sizes of processing units. 2) At the operating stage when the correct selection of values must be made for those variables which may be altered while the circuit is off-line but which are constant while the circuit is on-line, for instance the vortex finder diameter of the cyclones. 3) Under conditions of continuous operation when changes in the circuit feed or in the mechanical operation of the circuit may cause undesirable changes in the circuit product. The first two prob¬lems are concerned with optimization, that is, with making the best choice from many possibilities, and the last with control. Optimization Problems. The main purpose of a wet grinding circuit is to produce particles with a sizing distribution which will ensure optimum performance of the later chemical reaction or physical concentration circuits, because it is on the efficiency of these later circuits that the economics of the total mineral treatment process depend. The size distribution of the product from a grinding circuit may be affected by many variables, some concerned with the arrange¬ment and operation of the machines and some concerned with the ore characteristics, and it is important that these should be recognized in the design and operation of the circuit. For instance, in a simple ball mill-cyclone circuit the variables which affect the product size even for constant feed characteristics are shown in Fig. 19. In addition, the rate, size, hardness, and mineralogical composition of the feed will affect the performance of the circuit and the product size. When the circuit contains more than one mill, both the number of choices available and the complexity of the problem are increased consider¬ably. Consequently, one type of problem which is to be solved if an optimum product is to be obtained from an existing grinding circuit is that of circuit design, and this includes arrangement of processing units and selection of those operating conditions which cannot be altered when the circuit is operating, such as mill speed and cyclone inlet and outlet dimensions. The other major problem in optimization is encountered at the design stage when a choice must be made about the type, number, and size of processing units. For instance, for a typical high-capacity size reduction operation, should two circuits each containing one rod mill and two ball mills or three circuits each containing one rod mill and one ball mill be used? Alternatively, should ball mills only be used and if so, how many, what size, and what arrangement?

Jan 1, 1985

By S. J. Sindeband

Prior to discussing the metallurgy of sintered chromium borides, it is pertinent to outline some of the reasoning behind this investigation and the purposes underlying the work. This study was initiated as an aproach to the ubiquitous problem of a material for service at high temperatures under oxidizing atmospheres, and it was undertaken with a view to raising the 1500°F (816°C) ceiling to 2000°F (1093°C) or better. For the reason that no small, but rather a major, lifting of the high temperature working limit was being attempted, it was felt appropriate that a completely new approach be taken to this problem. A summary of the thinking behind this approach was published recently by Schwarzkopf.' In briefest terms, it was postulated that the following requirements could be set up for a material which would have high strength at high temperatures. 1. The individual crystals of the material must exhibit high strength interatomic bonds. This automatically leads to consideration of highly refractory materials, since their high energy requirements for melting are related to the strength of their atom-to-atom bonds. 2. On the polycrystalline basis, high boundary strength, superimposed on the above consideration, would also be a necessity. Since this implies control of boundary conditions, the powder metallurgy approach would hold considerable promise. Such materials actually had been fabricated for a number of years, and the cemented carbide is the best example of these. Here a highly refractory crystal was carefully bonded and resulted in a material of extremely high strength. That this strength was maintained at high temperature is exhibited by the ability of the cemented carbide tool to hold an edge for extended periods of heavy service. Nowick and Machlin2,3 have analytically approached the problem of creep and stress-rupture properties at high temperature and developed procedures whereby these properties can be approximately predicted from the room temperature physical constants of a material. The most important single constant in the provision of high temperature strength and creep resistance is shown to be the Modulus of Rigidity. On this basis, they proposed that a fertile field for investigation would be that of materials similar to cemented carbides, which have Moduli of Rigidity that are among the highest recorded. The cemented carbide, however, does not have good corrosion resistance in oxidizing atmospheres and without protection could not be used in gas turbines and similar pieces of equipment. It would be necessary then to attempt the fabrication of an allied material based upon a hard crystal which had good corrosion resistance as well. It was upon these premises that the subject study was undertaken and at an early stage it was sponsored by the U.S. Navy, Office of Naval Research. Since then, it has been carried on under contract with this agency. Chromium boride provided a logical starting point for such research, since it was relatively hard, exhibited good corrosion resistance, and, in addition, was commercially available, since it had found application in hard-surfacing alloys with iron and nickel. That chromium boride did not provide a material that met the ultimate aim of the study results from factors which are subsequently discussed. This, however, does not detract from the basis on which the study was conceived, nor from the value of reporting the results which follow. Chromium Boride While work on chromium boride proper dates back to Moissan,4 there has been a dearth of literature on borides since 1906. Subsequent to Moissan, principal investigators of chromium boride were Tucker and Moody,5 Wede-kind and Fetzer,6 du Jassoneix,7,8,9 and Andrieux." These investigators were generally limited to studies of methods of producing chromium boride and detennining its properties. Some study, however, was devoted to the chromium-boron system by du Jassoneix,7 who did this chemically and metal-lographically. This system is not amenable to normal methods of analysis by virtue of the refractory nature of the alloys involved, and the difficulties of measurement and control of temperature conditions in their range. Dilatometric apparatus is nonexistent for operation at these temperatures. Du Jassoneix made use of apparent chemical differences between two phases observed under the microscope and reported the existence of two definite compounds, namely: Cr3B2 and CrB. These two compounds, he reported, had quite similar chemical characteristics, but were sufficiently different to enable him to separate them. The easiest method for producing chromium boride is apparently the thermite process, first applied by Wede-

Jan 1, 1950

By F. A. Forward, J. Halpern

A new process for extracting uranium from ores with carbonate solutions is described. Leaching is carried out under oxygen pressure to ensure that all the uranium is converted to the soluble hexavalent state. By this method), alkaline leaching can be used successfully to treat a greater variety of ores, including pitchblende ores, than has been possible in the past. The advantages of carbonate leaching over conventional acid leaching processes are enhanced further by a new method which has been developed for recovering uranium from basic leach solutions. This is achieved by reducing the uranium to the tetravalent state with hydrogen in the presence of a suitable catalyst. A high grade uranium oxide product is precipitated directly from the leach solutions. Vanadium oxide also can be precipitated by this method. The chemistry of the leaching and precipitation reactions are discussed, and laboratory results are presented which illustrate the applicability of the process and describe the variables affecting leaching and precipitation rates, recoveries, and reagent consumption. THE extractive metallurgy of uranium is influenced by a number of special considerations which generally do not arise in connection with the treatment of the more common base metal ores. Perhaps foremost among these is the very low uranium content of most of the ores which are encountered today, usually only a few tenths of one percent. A further difficulty is presented by the fact that the uranium often occurs in such a form that it cannot be concentrated efficiently by gravity or flotation methods. In these and other important respects, there is evident some degree of parallelism between the extractive metallurgy of uranium and that of gold and, as in the latter case, it has generally been found that uranium ores can best be treated directly by selective leaching methods. It is readily evident that this parallel does not extend to the chemical properties of the two metals. Unlike gold, which is easily reduced to metallic form, uranium is highly reactive. It tends to occur as oxides, silicates, or salts. Two ores are of predominant importance as commercial sources of this metal: pitchblende which contains uranium as the oxide, U3O51 and carnotite in which the uranium is present as a complex salt with vanadium, K2O-2UCV3V2O5-3H2O. These ores may vary widely in respect to the nature of their gangue constituents. Some are largely siliceous in composition, while others consist mainly of calcite. Sometimes substantial amounts of pyrite or of organic materials are present and these may lead to specific problems in treating the ore. Further complications may be introduced by the presence of other metal values such as gold, copper, cobalt, or vanadium whose re- covery has to be considered along with that of the uranium, or whose separation from uranium presents particular difficulty. In general, there are two main processes for recovering uranium in common use today.'.2 One of these employs an acid solution such as dilute sulphuric acid to extract the uranium from the ore. A suitable oxidizing agent such as MnO, or NaNO, is sometimes added if the uranium in the ore is in a partially reduced state. The uranium dissolves as a uranyl sulphate salt and can be precipitated subsequently by neutralization or other suitable treatment of the solution. The second process employs an alkaline leaching solution, usually containing sodium carbonate. The uranium, which must be in the hexavalent state, is dissolved as a complex uranyl tricarbonate salt, and then is precipitated either by neutralizing the solution with acid or by adding an excess of sodium hydroxide. The latter method has the advantage of permitting the solutions to be recycled, since the carbonate is not destroyed. This is essential if the process is to be economical, particularly with low grade ores. With each of these processes, there are associated a number of advantages and disadvantages and the choice between using acid or carbonate leaching is generally determined by the nature of the ore to be treated. In the past, more ores appear to have been amenable to acid leaching than to carbonate leaching and the former process correspondingly has found wider application. With most ores, acid leaching has been found to operate fairly efficiently and to yield high recoveries. One of the main disadvantages has been that large amounts of impurities, such as iron and aluminum, sometimes are taken into solution along with the uranium. This may give rise to a high reagent consumption and to difficulties in separating a pure uranium product. Excessive reagent consumption in the acid leach process also may result

Jan 1, 1955

By Colin Baker, G. C. Smith

Investigation of the initiation and propagation of ductile failure in OFHC copper was undertaken to determine the role of nonmetallic inclusions. The effect of inclusion initiated voids on the formation of the internal cavity and the final shear separation was studied by metallographic eranzination of strained test pieces. A strain anneal technique was used to enlarge the voids under uniaxial stress conditions to elinzinate triaxial stress effects. Measurements of void size us stress and strain were made to show the point at which void im'tiation begins and becomes an important factor in the deformation process. The work of separation of copper-cuprous oxide was determined to attempt to correlate the breakdown of the matrix inclusion interface with void initiation and propagation. The zloid shape and position relative to the tensile axis suggested an interface breakdown mechanisnz of initiation. Evidence is presented that shows a basic similarity between the central cavity propagation and the 45-deg shear portions of the failure. DUCTILE fracture has been studied by a number of workers1-lo and attention drawn to the importance of hard second phase particles in the initiation of the failure. Holes formed at the matrix-particle interface can elongate by plastic deformation and then subsequently expand sideways to link up and produce a major crack. This is usually observed first in the center of the macroscopically necked region of a test-piece where the hydrostatic stresses are at a maximum. As the crack spreads sideways towards the free surface of the specimen, well defined shear zones develop from the crack tip and the final separation is along a direction at approximately 45 deg to the stress axis. This shear failure may also be associated with voids formed adjacent to second phase particles. In this way a cup and cone type fracture is produced. The stage at which separation takes place between particles and the surrounding matrix has not been clearly identified. In addition, although researchers have dealt with anisotropy of tensile behavior" as a result of material fabrication variables, not much is known about the microstructural features of aniso-tropic behavior. In the present work evidence on these points is presented in relation to the behavior of copper containing second phase particles of cuprous oxide. I. MATERIALS AND PROCEDURES EMPLOYED The material used was +-in. diam or 2-in. sq cold-drawn OFHC copper bar which contained 0.6 pct by volume of cuprous oxide inclusions. These ranged in COLIN BAKER, Junior Member AIME, formerly at -mnF of Metallurgy, University of Cambridge, Cambridge, England, is presently Research Scientist Reynolds Metals Co., Richrnand, Va. G. C. SMITH, Member AIME, is Senior Lecturer, Department of Metallurgy, University of Cambridge. Manuscript submitted June 20, 1967. IMD size from approximately 1 to 6 p in length and 1 to 4 p in width. The shape was generally slightly ovoid. Tensile tests were made on specimens having a gage length of 2.5 cm and diameter of 0.643 cm. Metallographic examination was carried out by sectioning deformed and fractured specimens; in addition fracture surfaces were examined optically and with a scanning electron microscope. Some measurements of the work of separation between copper and cuprous oxide were made, using a sessile drop technique which was a modification of that used by Kingery and umenick." The best metallographic results were obtained by using a vibratory polisher, which minimized smearing of the surface. 11. RESULTS A) Initial Experiments. Specimens from the +-in. diam rod were annealed for 2 hr at 650°C in uacuo, at which temperature complete recrystallization occurred without any change in the form of the inclusion. They were then fractured at temperatures from -190" to 600°C. Cup and cone fractures were obtained at all temperatures from -196" to 400°C. With increase in temperature there was, however, a continuous increase in the extent of the central transverse area and a corresponding decrease in the shear portion of the fracture. Above 400°C, the fractures became intergranular. Sections of specimens tested below 400°C revealed extensive small voids which were always associated with inclusions. However, the voids only reached dimensions greater than the inclusion size in the region of the macroscopic neck, where they were many times longer. Lateral expansion was found only near the fracture surface of the test pieces. As observed by Puttick, the voids were either (a) triangular holes initiated in the direction of the tensile axis and elon-

Jan 1, 1969

By Harry C. Swanson

TRANSPORTING concrete from mixer to forms has always been a problem. Twenty-five years ago this task was generally accomplished by means of wheelbarrow or concrete buggy. On large dam jobs, as the number of these projects increased, the gantry crane or highline came into use. Today several methods of handling concrete are employed on smaller surface construction jobs, for example, transit-mix trucks or dumpcrete trucks, which have crawler cranes with buckets for placing concrete into forms. In 1944, during early stages of developing Mather mine A shaft, several large underground concrete jobs were necessary. At this time the Cleveland-Cliffs Iron Co, purchased the first pump-crete machine, introduced by the Chain Belt Co. of Milwaukee. The machine was used to pour approximately 200 cu yd of concrete for a dam, or bulkhead, located 400 ft from the shaft. Concrete was mixed on surface, lowered down the shaft 1000 ft in a 2-cu yd bucket hung under one skip, spouted into the bowl of the pumpcrete machine from the bucket, and pumped directly into the forms. Since the day of the first pipeline concrete in 1944 to the present time, other equipment and other methods have been developed to permit transportation of concrete by pipeline through vertical and horizontal distances totaling 1 mile from mixer to forms. Much of the efficiency in present handling of underground concrete can be credited to the Bethlehem Cornwall mines, where concrete was transported through 6-in. pipe for great distances down an inclined shaft and along levels into forms.' During initial development of Mather mine B shaft, with concrete work under way on two or more levels at one time, the pneumatic concrete placer, Fig. 1, was selected as best adapted for underground concrete transportation. The 3/4-cu yd pneumatic placer is a small machine readily moved from one location in the mine to another. It can be equipped with two sets of mine car wheels, which will permit moving on regular mine tracks. It is therefore possible to send concrete through the pipe at great velocity; the pipeline is clean after each shot except for the film of cement adhering to the inside. With the proper slump in the concrete, it is possible to shoot concrete 2000 ft with this machine, using the mine supply of compressed air at 95 psi. This equipment was first used at Mather mine B shaft to concrete slusher drifts, Figs. 2 and 3, and finger raises located about 2000 ft from the shaft. In several instances there were bends into crosscuts and up vertical distances into the forms. For the first pours two placers were used. The first was located near the shaft where the concrete could be spouted into it from a 2-cu yd concrete bucket on the cage. The second was set on the side of the drift at a point approximately 1500 ft from the shaft. The concrete was shot directly into the second placer from the first unit and from the second machine directly into the forms. After completion of several pours with the two machines, a trial pour with only one placer located at the shaft proved that the second placer could be eliminated. Since then all pours have been successfully completed with only one placer underground. As production of iron ore from the mine increased and the development program expanded, use of the cage for handling mine supplies and concrete became a major problem. This brought about the first attempt at shooting concrete vertically down the shaft for 2600 ft. A 6-in. pipeline with victaulic couplings installed during shaft sinking was used for the trial. One placer was set on surface 250 ft from the collar of the shaft so concrete could be spouted directly into it from the mixer. This machine shot the concrete 250 ft horizontally on surface to the shaft, 2600 ft vertically down the shaft, and 100 ft horizontally into the second placer located near the rib of the shaft station or plat. The second machine shot the batch into the forms, about 2000 ft. Total distance horizontally and vertically was 4800 ft. The entire time cycle for a ¾-cu yd batch of concrete from the mixer on surface to the forms underground totaled about 5 min. During the past two years the two-placer method from the mixer on surface to the forms underground has proved a very efficient means of transporting underground concrete. Advantages of using pipeline concrete are as follows: 1—Interference with normal mining operation is eliminated. When the cage, skips, mine cars, or mine openings are used for transporting concrete and materials used for making concrete, mine operation suffers in one way or another.

Jan 1, 1955

By Aureal T. Cross

THIS paper is a review of the published literature on research in coal geology, principally exclusive of resource studies, which appeared or became available during 1950 and the latter part of 1949. This report is not to be construed as being complete. The papers referred to in the bibliography are those among many more, which were read either in full or in abstract. Undoubtedly other papers were published' which either escaped the author's notice or were not available to him. Those which were seen in abstract only (about one fourth of those listed) were not available in time for the inclusion of more than a notice. An outline of all papers listed in the bibliography has been arranged by subjects and reasonable subdivisions with some papers cited under more than one subject. Most papers are indexed according to the principal subject of discussion or research only as to an unusual or noteworthy section of the entire report. There will likely be some disagreement as to the quality or merit of some of the papers selected and the specialist may be supercritical of the outline or organization of papers in his field. It may be that attention has occasionally been drawn to papers reporting old information or conclusions of questionable value. Conferences and Meetings One of the best indications of the growing interest in coal geology problems in the United States is the increasing number of times this- field has been the focus of attention at conferences and meetings. Notable among these are the joint meeting of the Society of Economic Geologists and the Geological Society of America at El Paso, November 1949, at which the principal thesis was concerned with low rank carbonaceous fuel deposits, especially of western United States. Among the papers given which are already available were those presented by Barghoorn,16 Parry,180 Roe,153 and Parks.162 At the annual meeting of the Botanical Society of America in New York, December 1949, a joint meeting of the Paleobotanical and Microbiological Sections was held for which a symposium on Microbiology in Relation to the Geologic Accumulation of Organic Complexes was 'organized. Publication of the six papers presented by Ralph G. H. Siu, Elso S. Barghoorn, Irving Breger, Claude E. ZoBell, James M. Schopf, and A. C. Thayson is anticipated. At the regular meetings of the Paleobotanical Section at the same time, several other papers of interest reported on coal ball studies, partial coalification of petrified wood, and floras. In Chicago, April 1950, a symposium on Applied Paleobotany was held by the Society of Economic Paleontologists and Mineralogists in conjunction with the American Association of Petroleum Geologists. The five papers presented at this meeting dealt with the use of Paleozoic plant microfossils for stratigraphic work, J: M. Schopf, Devonian-Mississippian fossils of the black shales, Aureal T. Cross, Mesozoic plants of stratigraphic value, Th. Just, plant microfossils of the Tertiary, L. R. Wilson, and studies of the Brandon lignite, Elso S. Barghoorn. Early publication of these in the Journal of Paleontology is expected. The Nova Scotia Research Foundation and the Nova Scotia Dept. of Mines sponsored an excellent 3-day conference in June 1950, which, dealt with several aspects of coal geology. Papers on coal classification, P. A. Hacquenbard, structure and sedimentation problems in Nova Scotia, T. B. Haites, new techniques of thermal analysis, W. L. Whitehead, geochemical investigations of Nova Scotia coals, Irving Breger, the role of fossil plant spores in coal correlation and the stratigraphy of the coal-bearing strata of the Appalachian Region, Aureal T. Cross, were given. Some discussions of these papers by those in attendance were recorded, and the entire proceedings is being prepared for publication. In September 1950, an unusual 3-day field conference was held by the Ohio and West Virginia Geological Surveys under the sponsorship of the Coal Geology Committee. This study of the stratigraphy sedimentation, and nomenclature of the Upper Pennsylvanian and Permian coal-bearing strata of southeastern Ohio, southwestern Pennsylvania, and northern West Virginia .was augmented by two discussions on associated rocks (clays and shales) and stratigraphic nomenclature at Wheeling, and Morgantown, West Va. An extensive guidebook was prepared, and transcriptions of the Morgantown meeting were made. As a follow-up of the September field conference, a round-table discussion 'was held on this general topic at a special open meeting of the Coal Research Committee in conjunction with the November meeting of the Geological Society in Washington. Short prepared statements to invite discussion were given on each of several topics by L. M. Cline, Carl O.

Jan 1, 1952

By Charles M. Packer, Oleg D. Sherby, Roy H. Johnson

Originally round tensile specimens of a eutectic Zn-A1 alloy develop elliptical cross sections during superplastic deformation. This observation, coupled with a detailed study of the microstructure and preferred orieniation, suggests that crystallographic slip and continuous grain boundary migration or re-crystallization are important processes during super-plastic deformation. In spite of the extensive activity in superplasticity1-15 and the numerous explanations proposed, no single model has had universal acceptance. It has been established, however, that the general requirements for superplastic extension of two-phase alloys include an extremely fine, stabilized grain size of the order of a few microns, a temperature about equal to or greater than one-half the melting point, a critical range of strain rate, and a similarity in the mechanical strength of the major phases. The proposed models can perhaps best be characterized in terms of the important phenomena associated with them. These phenomena include: phase instability,1 diffusional creep by volume diffusion3 or grain boundary diffusion4,5 slip and continuous grain boundary migration or recrystalliza-tion,= grain boundary Sliding,7-9,13,14 and dislocation glide.'5 In this paper, experimental observations will be reported which support a model involving slip and continuous grain boundary migration or recrystalliza-tion. Specifically, a correlation will be made between this model and the development of elliptical cross sections as originally round specimens are superplas-tically deformed. For the most part, superplasticity studies have been conducted with eutectic or eutectoid alloys. Probably the most thoroughly studied material has been the monotectoid Zn-A1 alloy.1,2,6,12,13,15 No attention to the eutectic Zn-A1 alloy has previously been reported, and the results discussed in this paper represent part of a general study of the superplastic properties of this alloy. MATERIALS The alloys used in this investigation were prepared by melting appropriate quantities of 99.99+ pct A1 and 99.999 pct Zn in air, mixing, and pouring into a water- cooled stainless-steel mold. Wet-chemical analysis was conducted with each heat of alloy prepared, using the procedure of Fish and smith.16 The composition of the eutectic alloy was 95.1 wt pct Zn. Ingots about 2 in. thick were rolled to 0.4-in. plate at about 300°C with a reduction of 5 to 10 pct per pass. Specimens were machined from the plate with the tensile axis parallel to the rolling direction. The specimens were round, with 0.150-in.-diam, 1.25-in.-long gage length, and 0.25-in.-diam threaded grip sections. EXPERIMENTAL PROCEDURE Specimens were mounted inside a uniform-temperature quartz tube which was surrounded by a double elliptical radiant furnace with a 12-in.-long uniform-temperature hot zone and a low thermal capacity. The tube extended through the top and bottom of the furnace and permitted rapid quenching of the loaded specimens when quickly filled with cold water at the conclusion of the test. The quench precluded any effects on specimen microstructure from a normal, slow cool. Constant stress was applied to test specimens by suspending a load on a constant stress cam of the type described by Hopkin.17 The design of this cam permitted application of a constant stress for elongations up to 200 pct. For greater elongation, approximately constant stress conditions were maintained by systematically reducing the load manually. RESULTS As part of an investigation of the superplastic properties of the eutectic Zn-A1 alloy, evidence was obtained for the development of elliptically shaped cross sections as originally round specimens were extended. For example, after an elongation of about 100 pct, a round specimen with an initial diameter of 0.150 in. became elliptical with major and minor axis of 0.128 and 0.88 in., respectively. Photographs are presented to illustrate the ellipticity developed during superplastic deformation, Fig. 1. The specimen shown was deformed at a stress of 500 psi, at a temperature of 285°C, and a strain rate of 2.28 x 10-2 min-1. The strain-rate sensitivity exponent* was measured at *The strain-rate sensitivity exponent, m, is defined as d In o/d In c where o is the steady-state flow stress and E is the strain rate. this temperature and in the strain rate range 10"3 to 10-1 min-1 was found to be about 0.5. This value is typical of those observed with superplastic materials. The material studied exhibited negligible strain hardening during superplastic deformation, the creep rate remaining constant under constant stress and temper-

Jan 1, 1969

By J. E. Powell, F. H. Spedding

WHILE rare earths are reported to be widely distributed in nature and are not really rare," in practice, there are only a few minerals which are sufficiently rich in rare earths to serve as practical sources. Perhaps the best known of these is monazite which is a phosphate mineral containing rare earths and thorium. This mineral occurs as a dense brown sand in gravel beds and is particularly rich in the light rare earths of the cerium subgroup. This mineral is processed commercially for its thorium, cerium, and lanthanum content, and, consequently, furnishes rich concentrates from which neodymium, praseodymium, samarium, europium, and gadolinium may be obtained. Unfortunately, monazite is rather lean in rare earths heavier than gadolinium. A second mineral which is rich in the light rare earths is bastnasite, a fluoro-carbonate. Extensive deposits of this ore have been discovered in the western United States and have received considerable newspaper publicity in recent years. While bastnasite is very rich with respect to cerium, lanthanum, and neodymium, it contains even less heavy rare earths than does monazite. One of the better sources of heavy rare earths of the yttrium subgroup is gadolinite, a black silicate rock from which the rare-earth content can be extracted readily by acid leaching. It is obtained chiefly from Norway at the present time, although there are known deposits in the United States. Other sources of heavy rare earths include fergu-sonite, euxenite, and samarskite which are refractory tantalo-columbate ores. These minerals require caustic fusion or reduction to carbides with carbon before the rare-earth content can be extracted. All of the minerals which are rich in the heavy rare earths contain yttrium as a major constituent. After the rare earths have been extracted as a group from an ore by chemical means, it is generally convenient to precipitate them from acid media with oxalic acid in order to eliminate certain non-rare-earth impurities such as iron, beryllium, etc., which are usually present. The oxalate can then be readily ignited to R2O3. The oxide can be dissolved in acid and is the starting point for subsequent separation into the pure components. Perhaps the principal reason why the rare earths have not been studied as extensively as other elements of the periodic table, whose natural abundances are comparable, is that they are extremely difficult to separate from each other by the usual chemical means. Prior to 1945, the separation of one trivalent rare earth from another was a laborious process. All separations were based on repeated fractionation processes, i.e., fractional precipitation, fractional decomposition, fractional crystallization, etc. These processes were repeated from a few hundred to many thousands of times in order to obtain individual rare-earth salts of reasonable purity. Of course, mention should be made that, in the few cases where a rare earth could be oxidized or reduced to a valence state other than three, more conventional chemical means could be utilized to separate the oxidized or reduced ion from the other normally trivalent rare earths. The ionic states which deserve special mention are CeIV, SmII, Eu11, and Yb11. When it is possible to remove an element of the series efficiently, due to an optional valence state, its immediate neighbors also become easier to isolate. For example, binary mixtures of lanthanum and cerium, and praseodymium and cerium can be obtained by a relatively small number of fractional operations. The tetravalent state of cerium then allows the complete resolution of the binary mixtures by ordinary chemical means. Although the tetravalent state of cerium has been known for a long time, the divalent states of samarium, europium, and ytterbium were not used extensively in separations prior to 1930 because they are relatively unstable in aqueous media.'-" No attempt will be made to give a comprehensive review of the extensive literature dealing with the separation of rare earths. Rather, this paper will be confined to a review of those methods which have been developed at Iowa State College during recent years, and which have proved extraordinarily successful for the isolation of highly pure rare earths in quantity. It was obvious that, if pure rare earths were to become generally available, methods would have to be developed wherein the thousands of fractional operations made necessary by the similarity of rare-earth properties could be performed automatically. The development of chromatographic techniques and ion-exchange resins appeared to offer a mechanism by which this objective could be accomplished. A number of early attempts were made to separate rare earths by these means; for example, Russell and Pearce12 passed a mixture of rare earths through a cation-exchange column and reported

Jan 1, 1955

By A. M. Skov, L. F. Elkins

First response to large-scale water flooding in the fractured very low permeability Spraberry sand has led to a new unique cyclic operation. Capacity water injection is used to restore reservoir pressure. This is followed by marly months production without water irzjection and the cycle repeated. Expansion of the oil, rock and water during pressure decline expels part of the fluids but capillary forces hold much of the injected water in the rock. At least with reservoir pressure restored and with partial water flood development, field performance has proved this cyclic operation is capable of producing oil from the nzatrix rock at least 50 per cent faster and with lower water percentage than is imbibition of water at stable reservoir pressure. INTRODUCTION The Spraberry Field of West Texas presents unusual problems for both primary production and water flooding. Extensive interconnected vertical fractures in the fractional-md sandstone permitted recovery of oil on 160-acre well spacing, but they made capillary end effects dominant. Primary recovery by solution gas drive is less than 10 per cent of oil in place. The concept of displacement of oil from the sand matrix by capillary irnbibition of water has led to field techniques which promise greatly increased oil recovery. Free exchange of laboratory research, reservoir information and results of field pilot tests among the various companies has been very important in development of this technology. Five units covering a total of 170,000 acres have been formed for water flooding, and 10 other areas covering an additional 175,500 acres are in various stages of unitization. Part of the Driver Unit reaching fillup first has demonstrated very unusual waterflood behavior and indicated numerous operating problems that will develop within and among the various units. SPRABERRY ROCK AND PRIMARY PERFORMANCE The Spraberry, discovered in February, 1949, is a 1,000-ft section of sandstones, shales and limestones with two main oil productive members: a 10-15 ft sand near the top and a 10-15 ft sand near the base. In part of the field some thinner intermediate sands are oil productive, and others are water bearing. All sands have permeabilities of 1 md or less and porosities of 8-15 per cent. Ordinary core analysis and electric and radiation logs are ineffective in differentiating between oil productive and nonprcductive sands. Sands capable of containing producible oil are best identified by mercury injection capillary pressure measurement and, in some cases, by core water saturation. About 3,500 wells have been drilled in the 500,000-acre trend. Vertical fractures were observed in practically all Spraberry cores. Continuity and interconnection of fractures were confirmed by pressure interference among wells during early development.' Major fractures trend northeast-southwest as indicated by oriented cores and confirmed by five fluid injection tests, by analysis of the pressure transients observed during development,''' and by three interference tests in the Driver Unit Water Flood reported herein. Fracture spac- ing probably averages inches to a few feet. Spraberry wells typically produced 100-400 BOPD initially after hydrauLic fracture treatments. By 1962 oil production had declined to an average of 12 bbl/well/day, near the economic Limits of operation. Reservoir pressure had declined from 2,300 psi initially in the Upper Spraberry and 2,500 psi in the Lower Spraberry to 500-1,000 psi. Partial closing of the fractures with declining reservoir pressure is believed to be the cause of such low oil production rates at these relatively high reservoir pressures. Cumulative recovery of 208 million bbl of oil is 80 to 90 per cent of that recoverable by primary means. Performance of the entire reservoir is summarized in Fig. 1. IMBIBITION WATER FLOODING By 1952 reservoir performance indicated low primary recoveries. Most engineers, expecting serious channeling of injected fluids through the fractures, held little hope for secondary recovery. With its extensive background of research on the fundamentals of fluid flow within reservoir rocks, Atlantic's Research and Development Division on short notice in 1952 conceived that displacement of oil by capillary imbibition of water into the rock might significantly increase Spraberry recovery. Laboratory data reported by Brownscombe and Dyes scaled to probable reservoir conditions showed potential waterflood recovery equal to or greater than primary recovery with a 10-15 year flood life.= A pilot test using three 40-acre injection wells, one central producing well and 18 surrounding observation wells demonstrated technical feasibility of the process. Injection of 1.5 million bbl of water from November 1952 to August 1955 proved water entered the rock and displaced oil

By P. Hanniala, S. Harkki, B. Andersson

"The Outokumpu Flash Smelting Process was put into commercial operation in 1949. The original process was based on the use of preheated air. The adoption of oxygen enrichment of the flash smelting process air in 1971 was a crucial improvement, especially for smelter capacity. Oxygen use according to the Outokumpu method means selection of the most suitable oxygen content for the process air on the basis of the concentrate composition, capacity and matte grade. The furnace is not bound to only one fixed degree of process air oxygen enrichment; the oxygen content may vary from 21% to 100%.The degree of oxygen enrichment and the matte grade are the most important process parameters in reducing the operating costs of the whole system, including smelting, converting and acid production. Raising the matte grade allows more efficient use of the reaction heat in the smelting stage. Sulphur is recovered in a high-strength, constant-flow gas stream. The blowing time in the converters is reduced considerably.Oxygen enrichment and high matte grade improve the energy efficiency of the plant complex as long as the surplus heat from the system is not wasted. When the operating costs and investment expenditure of the total plant complex are taken into account, a reasonably high degree of oxygen enrichment (35-80°/0) is the best alternative in practice. Existing operating smelters have achieved single-stream thoughputs of over one million t/yr of concentrates.HistoryLocal capital and operating costs as well as lower energy consumption per unit of metal produced are now the crucial factors in determining the choice of process. In the late I940s, the same reasons prompted Outokumpu to develop the original flash smelting process in Finland. Between 1900 and 1940, very few technical changes in pyrometallurgical copper production took place. However, the commercial use of the flash smelting technique began a new era in metal extraction from sulphide concentrates. The characteristic feature of the method is that the concentrate is smelted in suspension: this offers the most favourable conditions for heat and mass transfer, and thus high reaction velocities. Outokumpu was the first to exploit the known advantages of flash smelting in a commercial installation at Harjavalta in 1949."

Jan 1, 1982

By Clarence G. Merritt

Stainless iron, as mild stainless steel is usually called, an alloy ranging from 11.50 to 15.00 per cent chromium with carbon under 0.12 per cent, has been considered to be not appreciably affected in respect to corrosion resistance by heat treatment. This metal, while one of the most widely used of the "straight" chromium stainless alloys, has failed in many applications because of insufficient information regarding heat treatment. These failures have been attributed to many possible causes from an inherent lack of resistance to corrosion and misapplication to mlysterious electrolytic effects or comparatively minute non metallic inclusions. Many uses for a steel with such excellent physical properties as are latent in this alloy have been lost on account of corrosion failures which could have been avoided by applying proper heat treatment. The alloy has not received the recognition or attention it should because of the fact that the details of its metallurgical behavior were not thoroughly understood. Response to heat treatment of the steel is well known as regards tensile properties,1-2 but its application has been for improvement in tensile characteristics or on account of machining considerations rather than for corrosion prevention. The potential user of this steel has not been clearly advised in papers, either technical or commercial, that have been written about it in the last 10 years or more, of particular features that are of prime importance in developing the excellent properties of this metal. The data to be presented in this paper are concerned mainly with the effect of heat treatment on the corrosion resistance, but a study has also been made of the concurrent microstructural changes and an explanation developed for the phenomena revealed, as well as a correlation to the other physical properties—hardness, tensile characteristics and impact strength. Stainless steel, largely used for cutlery purposes, an alloy of approximately 13.00 per cent Cr with 0.30 per cent C;, must be hardened and the surface prepared by polishing or pickling for best resistance to staining and corrosion. Monypenny reports that a loss of hardness is

Jan 1, 1932

By D. A. Whitley, R. P. Carone

"Shaft sinking by blind shaft drilling offers the mining industry a different approach to shaft construction. Benefits of blind drilling include less construction time and fewer safety problems while being economically competitive with conventional sinking methods under certain conditions in soft water-bearing formations.Technology and equipment are available today to drill blind shafts up to 20 feet in diameter and over 3,000 feet deep. This is an extension of proven techniques developed by private industry. To drill a 20-foot-diameter hole requires a specially manufactured hydraulic drill unit with a rated hoist of 2 million pounds and 500,000 ft/lb of torque.This brief paper will discuss the following aspects of blind shaft drilling:• benefits;• equipment f or drilling;• circulation methods;• circulation rates;• shaft alignment; and• progress rates to be expected.IntroductionShaft sinking by blind shaft drilling offers the mining industry an improved alternative method for shaft construction. The drilling of 20-foot -diameter shafts to depth s in excess of 3,000 feet can be done by utilizing equipment specifically designed and manufactured for this purpose. This equipment is a distant cousin, at best, to standard oil well drilling rigs. The actual configuration of these drilling machines will be discussed in a following section.To most people, the drilling of mine shafts may seem to be a new or at least recent technology. However, as early as the I920s, German and Dutch miners were drilling 24-foot diameter shafts for coal mines. What led ""miners to resort to drilling were ground conditions that still plague conventional shaft sinkers today: soft , water-bearing formations. The major drawback to these early attempts was a low penetration rate due to poor bit cleaning. This, however, was not the end of early shaft drilling activity. From 1936 to 1950, a number of mine shafts were drilled by the calyx or coring method. This method employs a bit with a large bore down the center. As it drills, it cuts a core. When 5 to 10 feet have been drilled, the core is removed. The disadvantages of this method were the time and work involved in breaking the core free, and it was inadequate in soft water-bearing ground."

Jan 1, 1981

By Wilson T. Lundy

The history of the production of sulfur from salt domes in Louisiana and Texas originated with the operations of the Union Sulphur CO. at Sulphur, La., followed by the Freeport Sulphur Co. at Bryanmound, Tex. Including the present development of Grande Ecaille by the Free-port Sulphur Co., which constitutes the tenth dome to be equipped for the production of sulfur in the Gulf Coast, less than 10 per cent of all the known domes have proved to contain sulfur in commercial quantities. The Grande Ecaille salt dome is in the tidal marsh of the Mississippi River delta in Plaquemines Parish, Louisiana. It is 10 miles southwest of the right descending bank of the river approximately 45 miles below the city of New Orleans and is within 4 miles of the shore line of the Gulf of Mexico. Terrain Characteristic of the delta region, the terrain consists of a low, flat, uninhabited area of marsh land intersected by many shallow lakes and bayous. With the exception of salt grasses, the region is devoid of vegetation and presents, in all directions, an unobstructed path to the vagaries of the winds. The general character of this area can be best visualized from the mechanism of its formation.' Offshore bars formed enclosing long narrow lagoons between the bar and the inner shore line, the lagoons communicating with the sea by means of tidal inlets. The water in the lagoons, being comparatively quiet, favored the deposition of fine debris composed of sand carried into the lagoons by action of the waves, tides and river-borne sediments from the land surface. As this material accumulated the bottoms of the lagoons were built up and eventually salt grasses obtained a foothold. The successions of growth and decay of these grasses formed a brown fibrous mat varying in depth from a few inches to several feet. This mat of vegetation virtually floats on an ooze that is several feet thick and is composed of alluvial sediments and very fine sand, and the whole is underlain by an unctuous clay. The decayed vegetation from the upper layer when dried has a peaty appearance.

Jan 1, 1934

By Cyril Wells, Robert F. Mehl

The purpose of this investigation was to prepare high-purity iron-carbon alloys, to determine as precisely as possible the A3(GOS), the Acm(SE), and the A1(PSK) transformation temperatures in the meta-stable iron-iron carbide system (up to 1.4 per cent C), and to study the influence of carefully controlled and constant rates of heating and cooling (1/8 to 2o C. per minute) upon the transformation temperature and upon the resulting microstructure. Apart from the general metallurgical importance attending the completion of such a task, it will be noted that highly precise transformation temperatures are of interest and importance in the thermodynamic study of this system1. Several determinations of the iron-iron carbide diagram have been made2. These suffer generally from the low purity of the alloys used, and from the application of methods of investigation of low sensitivity. Attempts to compare and weight the many data available in an effort to construct as accurate a diagram as possible have, in the main, been unsuccessful. In general, the wide differences in the results of different workers, the lack of complete analytical data, the omission of important details relating to the heat-treatment of the alloys, and the meager description of the methods of investigation applied, render such a process of averaging nearly meaningless. It may be seen from the summaries by Epstein2 and Körber and Oelsenl that variations as great as 30' C. exist in the reported data on the A3 temperatures in iron-carbon alloys. Fortunately, the Aa temperature in pure iron is now known to be 910" C. ± 1° C.3 The best recent work on iron-carbon alloys is that of Esser4 on hypoeutectoid steels and that of Sato5 on both hypoeutectoid and hyper-eutectoid steels. Esser made a dilatometric and magnetic study of iron and iron-carbon alloys containing 0.06 to 0.86 per cent C, for which a high purity is claimed. Presumably the electrolytic iron was reasonably pure, but the alloys contained considerable impurities, on the average totaling 0.35 per

Jan 1, 1937

By James K. Stanley

When a plastically deformed mctal is heated to a certain temperature, it undergoes a complete change in microstructure, the consequence of which is a marked alteration of mechanical properties such as hardness, strength and ductility. This change in microstructure occurs by re-crystallization and is a process involving the formation and growth of strain-free nuclei in the deformed matrix. These nuclei continue to form with time and grow until a new grain structure replaces the cold-worked structure. The formation of the nuclei in the deformed sheet of I per cent silicon ferrite is illustrated in Fig. I. If a recrystallization is conducted iso-thermally, a nleasurement of the fraction of the matrix recrystallized plotted against time leads to an isothermal recrystallization curve, first established in 1942.1 Such recrystallization curves are shown in Fig. 4. These curves show that the recrystallization of a cold-worked metal is best considered as a rate process rather than a time independent process such as is suggcsted by the improperly named and misleading recrystallization diagrams involving deformation, temperature, and grain size.2-5 Any study of the rate of recrystallization resolves itself into a study of the component rates of nucleation and growth. The previous paper1 provided an account of the determination of the rates of nucleation; N, the number of nuclei formed per unit time per unit of unre-crystallized area, and growth, G, the radial growth in unit time, in the recrystallization of silicon ferrite (I per cent Si). A method was developed to furnish experimental data for evaluating the effect of the recrystallization variables on N and G. The mathematical analysis for the derivation of the isothermal recrystallization curve was developed for a two-dimensional case (in thin sheets where the diameter of the recrystallized grain is greater than the sheet thickness). The present paper is a continuation of the study of recrystallization of I per cent silicon ferrite begun in 1942.1 In this work the effect of the factors deformation, temperature, grain size, composition, and recovery is studied with respect to the rates of liucleation and grouth. At the same time it is shown how the cooperation of N and G affects the isothermal recrystallization curve. The study of these variables on N and G has a definite bearing on the mechanism of recrystallization and should lay a foundation for a rational theory on the subject. Procedure The procedure used in these experiments was essentially the same as that described previously but in certain details some changes were made in the interests of speed. With few exceptions, the procedure is the same for the study of all the variables. Where the procedure is different from standard manipulations attention is called to this fact.

Jan 1, 1945

By George J. Clark

One of the major problems of operation in the anthracite industry is the disposal and handling of refuse—not because of its complexity but because of the quantity and type of material involved. It is the purpose of this paper to describe the kinds of refuse, methods of disposal, quantities of material to be handled and equipment used, together with present day application at specific operations in the Northern, Eastern Middle, Western Middle and Southern anthracite fields. For the purpose of this paper, refuse is considered to be any material that has no present or foreseeable future value and that must be handled and disposed of in the process of mining and preparing anthracite for market. Included in this classification are tunnel rock, mine rock, cleaner plant refuse, breaker refuse, old timber, ashes and silt. Disposal Methods Old timber is disposed of either by burning or by depositing on the refuse bank. Burning is preferred because timber on the refuse bank is a potential fire hazard. Ashes are deposited on a separate bank, mixed with other refuse, or flushed into old mine workings. In many places, the latter practice has caused considerable difficulty in preparation of coal mined from these areas. In depositing ashes on the surface, it is wise to refrain from mixing them with material containing coal as such contamination may prevent reclamation of the coal in future years. Tunnel work, mine rock, cleaner plant and breaker refuse compose the bulk of the material to be disposed of and handled, and are referred to collectively as "refuse" in this paper. Waste material in or adjacent to the coal beds may be removed at the working face wherever possible. This can be done in flat or lightly pitching beds, and in the Northern field represents a sizable percentage of this type of refuse handled. In the other three coal fields, this method of disposal cannot often be used. Refuse not removed at the working face is taken to the surface and generally is deposited on banks; occasionally it is crushed and flushed into old mine workings. Breaker refuse often is used for constructing berms for silt-retention basins. Disposal Problems The problem of handling refuse is different for each operation, owing to topographical conditions, quantity of material to be handled, anticipated requirements and types of equipment in use elsewhere at the operation. Preparation plants with limited refuse-disposal areas, excessive grades to reach the required elevation and long hauls to disposal areas not readily accessible require the selection of equipment to best meet the conditions. The quantity of refuse to be handled from a given amount of raw material varies considerably throughout the four fields. In mine coal, this varies from a low

Jan 1, 1949

By R. C. J. Goode, W. S. Rapson, W. R. Lawrie, L. W. P. Van Der Bosch

W. R. Lawrie (Member): As all the Institute members present are aware, the Ninth Commonwealth Mining and Metallurgical Congress was held in Great Britain from 3rd to 24th May, 1969. The Congress opened in London and the first week was devoted to technical sessions. Several receptions and other forms of entertainment were held in the evenings. The Congress was attended by some 1 200 delegates, many of whom were accompanied by their wives. As can be imagined a large organisation was needed to cater for all the whims of these delegates, to arrange accommodation and to provide transport for the technical trips, for the receptions and for the sight-seeing tours. Papers presented at the technical sessions proved of great interest. It was stimulating to hear of the developments taking place in the various branches of mining and metallurgy. The discussions which developed between men from so many different countries were thought provoking, led to new friendships and most of us made valuable contacts for the future. There were 141 papers presented in the one week. They were given in four separate halls with two, three and sometimes four papers coming up for discussion in each of the three sessions per day in each hall. Among the more interesting subjects were those concerning off-shore drilling and the facilities for production of petroleum and natural gas. Beach mining at Consolidated Diamond Mines drew a large audience. Automation, mechanization and other technical developments were well described and we were brought up to date with modern techniques of mineral prospecting. There was general discussion on the future of the mineral industries together with comments on the consumption and price trends of these metals and minerals. The future of uranium, of vital interest to us in South Africa, came in for much debate. During the first week, besides attending the technical sessions and seeing some of the sights of London, we were royally entertained. There was a banquet, attended by some I 400 people, where we were welcomed by Princess Alexandra, a reception by H.M. Government in the Banqueting House, and we were entertained at the House of Lords on a terrace overlooking the bustling and rather muddy river Thames. Then there was the visit to the Glyndebourne Opera. It is only in London that at 3 o'clock of an afternoon, nearly 1 000 people all togged up in dress suits and long dresses and carrying packets of sandwiches for supper could arrive at a station to catch a train, and not even cause a stir of interest or a raised eyebrow! This is the way we travelled 60 miles to the opera from London. Surely there can be no more picturesque or romantic a setting for an opera than this large 400-year old red brick manor house nestling in the green valleys of the unspoilt Sussex countryside. It was a glorious evening and though we returned somewhat late and tired, not one of us would have missed this experience. After the first week in London there was a choice of tours of one week to London and the Home Counties, Cornwall, a Geological tour to Scotland, South Wales, and North England. These were followed by further one week tours to Yorkshire and Lancashire, Scotland, West Country and Midlands, a Geological tour of Wales, and a Mining and Geological tour to Jurassic Iron Mines. We then returned to London for the final session and closing banquet. Post-congress tours were arranged to Europe and to Ireland. It was a wonderful Congress, a great experience, we made many friends and may there be many more congresses! . L. W. P. van den Bosch (Member): Much has already been said in appreciation of the excellent organisation and arrangements made for the delegates who attended this Congress. This report is confined to a brief discussion of the personal impressions gained from the papers presented and the technical visits attended. The papers covered a wide variety of subjects ranging from highly theoretical observations to practical descriptions of operations. There were three main themes, VIZ: 1. The attention given to research and its forceful application to practical operation. 2. A steady development in mechanization and automation leading to savings in manpower. 3. Improvization, modernization and adaptation of exisitng facilities. Thoughts on these themes can best be illustrated by comments on some of the industries visited. COAL The National Coal Board (N.C.B.) has been most progressive and all of us have heard of the Bevercotes Colliery with its completely automatic mining and coal handling equipment. This was not, unfortunately, on show but there is no doubt that this is one of the greatest advances in coal mining leading towards continuous production. Coal in Britain is fighting for its life against oil, North Sea gas, imported gases and nuclear power. Output is dropping, the less efficient collieries are closing down and efficiencies are steadily rising as the following tabulation indicates: In 1931: 1 million men produced 300 million tons of coal In 1947: 0.7 million men produced 230 million tons of coal In 1968: 0.4 million men produced 160 million tons of coal