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By Ross W. Smith, Salim Akhtar

INTRODUCTION Flotation of oxide minerals, particularly silicates, with cationic collectors differs substantially from classic flotation of sulfide minerals using sulfhydryl collectors. In the case of cationic flotation of these minerals, in general: 1. There is a greater importance of electrostatic interaction between mineral and collector. 2. Longer-chained collectors are required; as a rule at least 10 carbons must be present in the hydrocarbon chain. 3. Moderate concentration of collector is required. 4. Attachment between collector and mineral appears to be less strong even though the system may exhibit a large contact angle. 5. Required conditioning time before flotation is usually relatively short. 6. There is sensitivity to slimes. 7. The collectors usually possess frothing power. 8. Selectivity of cationic collectors for specific minerals is usually poor. Almost any of the oxide and silicate minerals may be floated using cationic collectors. However, specific response to these collectors varies with mineral type. The following classes and minerals are

Jan 1, 1976

By Edward C. Mack, Ronald G. Stovash

INTRODUCTION Ronald G. Stovash General Superintendent Consolidation Coal Company Shoemaker Mine Moundsville, West Virginia Consolidation Coal Company's Shoemaker Mine is located in West Virginia's Northern panhandle near Wheeling, adjacent to the Ohio River. Steam coal is mined from the Pittsburgh #8 seam which averages 65 in. in thickness. The 10 in. of draw rock directly above the coal is taken on advancement, but remains when longwalled. The top consists of up to 12 in. of lower quality coal, and immediately over that, up to 10 ft of multi layered shales, coal, and bone. The cover ranges from 300 to 800 ft of shales and limestone. Two longwalls operate in the mine. The first longwall started in April 1975 and is retreating its eighth panel. This wall was the first to use shield supports in the United States. The second longwall started in September 1979 and is currently retreating its first 7,000 ft panel.

Jan 1, 1981

By Benjamin C. Koskiniemi

Introduction When evaluating any ore body, one of the first questions concerns the ore reserves. In the case of an open pit mine, this is not possible to answer reliably until the ultimate (final) pit limits have been established. Techniques used in designing an ultimate pit are classed as (1) manual, (2) computer, and (3) combined manual-computer. This chapter will describe how manual techniques can be utilized in designing an ulti¬mate pit Certain economic and design criteria must be established before the actual design begins In order to begin designing an ultimate pit, it will be assumed that the engineer already has the following data available (1) vertical sections, (2) horizontal sections for each level, (3) stripping curve, (4) bench height, (5) bank slope angle between levels, (6) level berm width, (7) roadway width, (8) pit slope angles at ultimate pit limits (estimated average including roads and ramps and between roads and ramps), and (9) minimum width of pit bottom

Jan 1, 1979

By W. W. Wen, S. C. Sun

The objective of this investigation was to study the electrokinetic behavior of oxidized coals to indicate the feasibility of separating oxidized coals from ash materials and pyrite by an amine flotation process. The effects of soluble inorganic electrolyte products from coal oxidation were also examined. The electro-kinetic behavior of coals of different rank and degree of oxidation were studied by an electro-phoreses technique. Generally, increasing the oxidation increases the negative value of the zeta-potential and decreases the isoelectric point (IEP). Flotation results indicate that the flotation recovery increases with decreasing pH values, and that with ferrous sulfate at pH 6.5 the flotation recovery is optimum while the pyrite is depressed. Comparing the flotation results with the electokinetic results verified that in the presence of ferrous sulfate, the maximum hydro-phobicity of oxidized coal with dodecylamine and frother is near its IEP, which is around pH 6.5 to 7.0.

Jan 1, 1978

By T. M. Olson, B. G. Lewis

The adsorption of selenite on bentonite, kaolinite, and an iron--coated illite was studied to determine the importance of several soil variables on selenite removal from uranium mill tailings pond seepage. Adsorption capacities of the clays of selenite varied by orders of magnitude and decreased in the following order: illite > kaolinite > bentonite. Maximum adsorption of selenite on the clays was observed between pH 2.5 and the clay's zero-point-of-charge. Substantially less selenite was adsorbed on these clays in actual mill tailings solutions than in model systems. Virtually all selenite was removed from the tailings at pH 7. The principal removal mechanism at this pH is suggested to be the adsorption of selenite on freshly precipitated metal hydroxides. It is concluded that selenite can be removed from tailings seepage in soil or in clay liners in the presence of kaolinitic clays and iron oxides. Selenite would also tend to be removed from tailings seepage in calcareous strata.

Jan 1, 1984

By A. E. Himebaugh

A SEMI-AUTOMATIC hoist has been installed at the 2966 station of the Copper Queen mine, since neither the bulk of supplies handled nor the number of men engaged in developing the 3100 level justifies the constant attendance of two cagers and two hoist operators. The connecting service shaft between the 2966 and 3100 levels is actually a two-compartment raise. Below the 2966 level the shaft is 7x13 ft-slightly oversize to permit downcast ventilation and provide space for the 16-conductor flexible control cable. The raise extends 52 ft above the 2966 level for a 3-ft diam head sheave. Another 3-ft sheave is mounted as an idler between the hoist and head sheave, eliminating the need for an inclined cableway. The main compartment serves as a cageway; the other compartment provides a manway and space for service lines. Since vertical travel is limited, rope speed is of minor importance.

Jan 5, 1958

By R. K. Iyengar, G. C. Duderstadt

An investigation was carried out to determine the effect of Ca-Si additions on the dissolved oxygen content of liquid steel. An apparent equilibrium was reached after holding the melt for some time when the total oxygen content of the melt was identical with the dissolved oxygen. Results of the investigation show that for deoxidation with silicon and manganese the apparent equilibrium is reached after 8 to 12 min and the oxygen content of the melt is in good agreement with values reported in the literature for similar steel compositions. Ca-Si additions decrease the dissolved oxygen content appreciably below that obtained with Si-Mn deoxidation. It is postulated that some calcium dissolved in liquid steel at the time of its vaporization combines with dissolved oxygen to form CaO which then fluxes the manganese silicate, thereby lowering the activity of the deoxidation products. Slow flotation of calcium silicate inclusions is attributed to their slower growth rate. An addition of aluminum (to yield <0.005 pct Alsol) prior to introduction of Ca-Si improves the kinetics of removal of resulting deoxidation products. WITH the introduction of continuous casting of billets and blooms, application of calcium-containing alloys for deoxidation purposes has gained new interest. More so than in conventional ingot casting, the degree of control of the dissolved and total oxygen contents in steel for continuous casting can determine the success or failure of the operation because these affect both surface (pinholes) and internal quality of the billets. Of the available deoxidizers, only silicon has so far found wide application. However, in the typical range of application (0.20/0.30 pct Si), it is too weak a deoxidizer to suppress pinhole frequency to the level desirable for all but the least demanding appli-cations (approximately 10/50 pinholes per sq ft, de-pending on degree of subsequent reduction). On the other hand, Vincent&apos; has shown that it requires ap-proximately 0.007/0.008 pct A1 soluble in the steel to suppress pinhole formation to <10 pinholes per sq ft. In view of the difficulty of consistently maintaining this aluminum level and the associated problem of tundish nozzle constriction2 when aluminum is added to the ladle, steelmakers have searched for other de-oxidation alloys to circumvent the problem. Calcium-containing alloys offer such a possibility. When employed as a partial substitute for silicon, CaSi (30 pct Ca/60 pct Si) additions are reported to yield adequate pinhole contro1.3 Its use simultaneously preserves the "fluidity" of the steel and ensures good castability.4 Oxygen contents as low as 0.007 to 0.014 pct have been obtained with CaSi additions varying between 3 to 8 lb per ton.5-7 Due to the high degree of vaporization of calcium at the temperature of molten steel and the different methods used to introduce it into the liquid, the reported results show considerable scatter and lack of consistency. According to several investiga- deoxidation tors, the flotation characteristics of CaSi products are not as favorable as those found for alumi-num but are similar to those established for silicates with low alumina content.678 Thus, cleanliness ratings were improved when CaSi was replaced by a com-bination of Si + A1 followed by CaMnSi additions.5,9 In view of the lack of a comprehensive description of the deoxidation potential of CaSi alloys, a study was undertaken of the effect of CaSi additions on oxygen content in comparison to Si + Mn and Si + Mn + Al. The alloy chosen for this study contained 63 pct Si, 32 pct Ca, <3 pct Al, and 2.5 pct Fe. EXPERIMENTAL PROCEDURE The experiments were divided in three groups as shown in Table I. In each group, the CaSi addition was increased successively while maintaining the melt composition constant. The experiments in the third group were designed to determine the effect of prede-oxidation with small amounts of aluminum. Ingot iron, containing 0.02 to 0.03 pct C and 0.03 to 0.05 pct Mn, was melted in a 100-lb magnesia crucible. Pig iron was added to the melt under air in an induction furnace to attain the desired carbon content of 0.05 to 0.15 pct and thereby control the initial oxygen content of the melt between 0.04 and 0.015 pct. After removal of the slag, the furnace was sealed and argon was introduced through an opening in the graphite cover, Fig. 1. When the desired temperature was reached, electrolytic manganese was added to the melt and pin samples were taken (7 mm silica tubing) to establish initial oxygen content. To obtain maximum efficiency, CaSi and metallic silicon were introduced in a sealed steel pipe into the bath. This procedure assured minimum loss of calcium through vaporization as the deoxidants were always released at the bottom of the melt. Samples were taken at regular intervals and quenched in water. Temperature of the bath was measured with a Pt/Pt-10 pct Rh thermocouple. Portions of the pin sample were used for oxygen analysis; five determinations per sample were made to obtain average oxygen content. Oxygen analyses were made by the inert carrier gas fusion method with frequent cross checks with the vacuum fusion method.

Jan 1, 1970

By R. D. Groves, G. M. Potter, T. H. Jeffers

Experiments on ammonium carbonate leaching of native copper ores crushed to 1-inch size showed that in 30 days 55 percent of the copper was extracted from 0.7 percent copper conglomerate-type ore, and 85 percent was extracted from 2.1 percent copper amygdaloid-type ore. Cupric ions were not necessary to initiate leaching, and dilute solution containing only 2.2 grams per liter NH3 and 1.9 grams CO, was a suitable solvent. Copper was recovered from the leach liquor by solvent extraction and electrolysis of the strip solution. Cyclic testing of the leach-solvent extraction-electrolytic procedure is in progress on a 6-ton charge of minus 12-inch conglomerate-type ore assaying 2.2 percent copper. Oxygen is metered to the closed leach vessel as required by the leaching reaction. Initial copper dissolution was much slower than from minus 1-inch ore, but the leaching rate then markedly improved.

Jan 1, 1974

In high-volume mining operations, preventive maintenance programs that include standardized replacement wear parts can boost equipment availability and lower unit costs. Meeting the daily production targets depends heavily on keeping equipment working as close to full time as possible. This is the case at one medium-sized western surface coal operation producing approximately 2.7 million t/y (3 million stpy). The mine, located in western Wyoming, removes in excess of 8 million m3 (10.5 million cu yd) of highly abrasive overburden each year. According to the mine superintendent, "Abrasion at this mine is some of the worst I&apos;ve ever seen. Shovel points can wear out in little more than one shift. Dozer blades some- times last two weeks, but can wear out sooner. And ripper points really take a beating. If we can replace this fast-wearing iron with small pieces that go on quickly, we think that&apos;s the way to go."

Jan 3, 1978

We did install one 100% brick bottom in our #33 furnace which is basically the same design as Inland&apos;s furnace. We had previously embarked on a program of replacing all bottoms in our 300 ton furnaces every 2nd campaign which would be about every 2000 - 2300 heats. This had been necessitated by the effect of our rather high oxygen blowing rates and our desire to maintain a minimum bottom interval level. This meant that we were able to compare bottom performance of various types of materials on a realistic bas is. Since furnace availability is critical to our operation, we are organized to employ every piece of equipment and all the necessary man power to complete a rebuild in the shortest possible time. A 300 ton furnace requires 7.5 days to rebuild. If a new bottom is to be installed an additional 2.5 days is added - 24 hours for tear out and 36 hours for bricklaying - making a total of 10 days.

Jan 1, 1972

By R. Adair, W. R. Hudspeth, W. T. McDaniel

D. C. Ralston (U. S. Bureau of Mines, Washington, D. C.)—Flake mica can be beneficiated by a wider variety of methods than almost any known mineral. However, most of these methods are not recorded. It is therefore a pleasure to find a written record of another application of a new method of beneficiation. The fact that the Humphries spiral gives a concentrate containing the mica flake mixed with the slime and small grit has forced a further separation by screening, in which the finest of the flake will pass through. Fine flake has its uses and if mixed with slimes or clay is recoverable reliably only by flotation. This loss must be charged against the spiral. On the other hand, the new muscovite mica paper recently brought out by General Electric Co., can be made only with grit free flake larger than 50 mesh. It is of interest to learn from the authors whether the washing of the flake on a 50 mesh screen can take the grit content below 1 pct, to meet this new demand. Must another degritting process be used on the flake, like electrostatic treatment of the dried material? H. D. Snedden (The Humphreys Investment Co., Denver)—When recovering mica in the spiral, gangue particles coarser than 16 mesh can be rejected more readily when operating at a high pulp density (35 pct solids) or a low volume rate (—15 gpm). A launder screen as used by Lehigh Coal Co. has proved very satisfactory to partially dewater the mica concentrate and to reject any gangue slimes it may contain. Under certain conditions mica cleaning in the spiral can be improved by reagentizing the spiral feed, the object being to insure wetting of the gangue minerals and to agglomerate or prevent wetting of the mica particles. R. B. Adair (authors&apos; reply)—The grit content of the mica recovered on Humphreys Spirals varies with the fineness of the launder screen used to clean the mica product. Normal practice has been the use of 80 mesh screen, which will produce a mica product of acceptable grit content for most purposes. By using a 50 mesh screen, and increasing the screen area per ton of feed, it is felt by the authors that a mica of 99 pct plus purity could be obtained. This statement would hold only in the case of mica feed which was well de-laminated, and free from "clay balls," (semikaclinized feldspar), which tends to follow the mica through the process.

Jan 1, 1952

By D. R. Parrish, V. J. Berry

Reverse combustion is one thermal method of recovering hydrocarbons from porous undergrortnd formations containing oil or tar. In applying this method, air is introduced via an injection well and the mixture of air and hydrocarbons is ignited in the production well. A combustion zone then recedes toward the injection well, counter-current to the air flow. If the combustion-zorie temperature is sufficiently high, the oil or tar irz place is distilled arid cracked. The hydrocnrhon flows as a vapor to the production well and subsequently is condensed at the surface. Maximum temperature and velocity of morcment arc the two dependent variubles defining the progress of the combustion zone. A theoretical analysis has been made of heat flow in the reverse-comhustion process as.srtrning linear flow in a homogeneous system. The differentirll equatious, which include the oxygen-hydrocar-hon rctrction rate, have been solved numerically. Rescllts indicate that the maximum temperature reached and he cornbustion-zone relocity both increase with an incrense in air-injection rote. Heat loss to surroundings has little effect on the maximum comhustion-zone ten?perature achieved, but it is reflected in a reduced com-bustion-zorzc, velocity. It is also predicted that an increase in the oxygen-hydrocarbon reaction rate results in a reduction in the maximum temperature reached. The calculated results are in agreement with results from reverse-combustion experinments using samples of a tar sand. INTRODUCTION Reverse combustion is one thermal method of recovering hydrocarbons from porous underground forma-tions containing oil or tar.&apos; In applying this method, air is introduced into the underground formation via an in-jection well. In one or more adjacent production wells, the mixture of air and hydrocarbons is ignited at the sand face. The combustion zone thus formed recedes toward the injection well, counter-current to the air flow (Fig. 1). Under certain conditions, the rate of recession of the combustion zone is sufficiently rapid that only a small fraction of the hydrocarbon at any point in the formation is consumed by combustion. If the combustion-zone temperature is sufficiently high, the uncon-sumed oil or tar in place may be partially distilled or cracked. This material then can flow as a vapor with the gaseous combustion products to the production well and subsequently be condensed at the surface. Under other conditions, the combustion-zone temperature may be so low that little, if any, of the hydrocarbon is vaporized. Nevertheless, the viscosity of the hydrocarbon may be reduced enough so that other recovery methods subsequently can be applied to produce oil at attractive rates. The maximum temperature achieved in the combustion zone and the velocity of movement are the two dependent variables defining the progress of the combustion zone. Presumably. knowledge of these factors will permit estimates to be made of oil recovery efficiency, oil and gas producticn rates and average amount of injection air required per barrel of oil recovered. The maximum temperature and combustion-zone vclocity are, in turn, expected to be dependent upon air flow rate, heat losses to surrounding formations, physical properties of the rock and the hydrocarbon. the extent to which the porous formation is saturated with the hydrocarbon, the rate of oxidation of the hydrocarbon and the amount of heat released. It is desirable that the relative importance of each of these factors be understood so that the results of laboratory tests of oil recovery by reverse combustion can be properly interpreted in terms of expected field performance. It is particularly desirable to assess the importance of heat lcsses because these are difficult to control or eliminate

By H. B. Charmbury, D. R. Mitchell

Recovery of fine coal from solids reporting to wash water has become increasingly important. These solids range from about 28 mesh to 0 and constitute 3 to 7 pct by weight of the feed tonnage to a coal preparation plant. In the past they were usually wasted to streams, but present clean stream legislation makes it necessary to remove them from the wash water and, if economically possible, recover the associated coal as a saleable product. Where these solids are not wasted to streams or to slime ponds they are recovered by thickeners and filters. Since filter cake is often high in ash or sulfur, or both, it is not readily marketed or used. Common practice is to mix it with other plant products, generally slack or cleaned run-of-mine. Uniform blending is difficult, not only because the impurity is usually high, but also because during certain operating periods it is possible to get a concentration of this filter cake in a railroad or barge.

Jan 2, 1959

Thirteen Directors, the Secretary, and seven guests were present. The formation of a Philadelphia. Section was discussed. The proposed rules for admission of additional societies to membership in Engineering Council were approved. An invitation was extended to the American Institute of Metals to become affiliated with the Institute upon the general terms proposed in the report of the Joint Conference Committee of the American Institute of Metals and the American Institute of Mining Engineers. The request of the Iron and. Steel Committee to hold a meeting in Milwaukee during the week of Oct. 7, 1918, was granted. The reports of the Treasurer and Finance Committee were accepted. A change in the By-laws was proposed. Thirty Members, two Associates, and four Junior Members were elected. The status of two Junior Members was changed to Member. The resignations of seven members were accepted. The clues of two members who are on active duty at the front were remitted. One member was reinstated to membership.

Jan 5, 1918

By Garth M. Crosby, F. McIntosh Galbraith, Bronson Stringham

THE Hercules mine is located in the northeastern section of the Coeur d&apos;Alene district, approximately 1 1/2 miles north of the town of Burke, Idaho. Surface indications of the ore deposit were first discovered in 1886, but regular mine production was not started until 1902 and was continuous until April 1925, when the known ore had been extracted. Incomplete records show that from 1912 until operations were suspended the mine produced 2 1/2 million tons of ore containing 9.4 pct lead and 7.7 oz of silver per ton, together with an estimated 2 pct zinc, 0.3 pct copper, and 20 pct iron. This operation was the first in. a series of mining enterprises culminating in October 1947 with the consolidation of Day Mines, Inc. In the same year it was decided to unwater the levels below the collar of the Hercules shaft in the hope of finding some indication of a recurrence of ore. The unwatering operation has been described in a. previous paper.&apos; The initial exploration, following recapture of the workings, showed sufficient promise to warrant a detailed study of the mineralogy with modern techniques. The general geo1ogy of the Coeur d&apos;Alene district, including a detailed description of the rock types encountered, has been comprehensively treated by Ransome and Calkins&apos; in their classic paper, and only local background description, therefore, is felt to be appropriate here. The Hercules deposit transects a portion of the trough of a broad south-trending synclinorium which has been greatly complicated by faulting. More locally, it lies within a block of ground bounded on the east; by the O&apos;Neil Gulch fault, a steep north-south overthrust of considerable magnitude, and on the west by a monzonite stock, the outcrop of which is 1/2 mile or more wide and 5 miles long. The country rock is composed of thin to medium-bedded argillites and argillaceous quartz-ites of the Prichard and Burke formations, the oldest members of the Pre-cambrian Belt Series of sediments in the area, believed to be of Algonkian age. The contact between them is a conformable gradation. The argillite is colored gray to tannish-gray and is fine-grained, compact, and generally massive in structure. Under the microscope the unaltered argillite is seen to be composed principally of anhedral quartz and a few feldspar grains which were at one time presumably partly rounded sand grains, but as a result of recrystallization and cementation by silica, the interstices are now almost obliterated and quartz grains show crenulate boundaries. The sizes of these crystals vary from 0.5 mm down to 0.1 mm in greatest dimension. In all specimens sericite comprises 10 to 20 pct of the rock and is present abundantly between most of the grains as flakes or shreds which vary considerably in size. Sometimes they form a fine felt-like mat or aggregate, and sometimes flakes are seen which appear to be good muscovite. In some specimens, separated rhombic-shaped carbonate grains are abundant, and in some instances these have been changed to sericite. Mining operations to date have explored the Hercules vein to a maximum vertical depth of 3600 ft below its outcrop, and along a maximum strike-length of 3600 ft on certain of the lower mine levels. The main orebody is irregular in outline, extending over a variable strike-length of 400 to 1500 ft; and it is intersected by a strong transverse fault that has been traced from the surface to the bottom level. This has been named the Hercules fault, and apart from the vein itself, it is the most prominent structural feature in the mine. There is good evidence that it existed prior to the introduction of ore solutions and may have influenced ore deposition, but it was also the locus of important post-ore displacement and shows a progressive right-handed horizontal component reaching 200 ft on the deeper levels. Its vertical component is not definitely known but may be considerably greater. The fault strikes 20° N to 50° E and dips westerly at angles of 70" to 45", flattening in dip where it crosses the original orebody from east to west between 1000 and 1600 ft below the surface. At about 3000 ft in depth the Hercules fault is joined by a vertical fault of similar strike, and the major post-ore dis-placement below their junction is taken up along this vertical branch of the structure, now called the Mercury fault. Recent work has been concentrated in this vicinity. Another structural feature of special geologic interest, though of little economic importance, is the occurrence of a porphyritic dike in this area. This lies a short distance above the Hercules fault, essentially parallel to it, and is 5 to 15 ft in thickness. It appears at first glance to cut the mineralization, suggesting push-apart relationship, but small stringers of the vein minerals have been observed to penetrate the dike for a matter of inches at several points. The dike is thought to be related to the monzonite intrusion. A vertical longitudinal projection of the mine is shown in Fig. 1, which illustrates most of the features discussed above. The Hercules vein was deposited along the course of a strong, persistent shear zone that now appears as a braided network of gouge seams running through more or less crushed and shattered country rock. It strikes 70° N to 80° W and dips southerly at an average of 75". Barren parts of the structure vary in width from less than 1 ft to more than 15 ft. The width of mineralized segments may be double that. Although the evidence is not conclusive, pre-mineral, normal movement along the zone may be 1000 or 1500 ft. The horizontal component is unknown. Post-ore movement appears to have been

Jan 1, 1954

By R. A. Falconer, H. L. Lovell

With increasing requirements for fine coal cleaning and the many advantages of the cyclone operating on a coal-water slurry, a series of studies designed to achieve an enhanced understanding of the cyclone parameters has been pursued. The purpose of this investigation has been to vary a series of dimensional parameters with an experimental 15-in. diam hydro-cyclone, and to determine the relative effects of parameter variations upon ash and sulfur removal. This study of the hydrocyclone is a continuation of a series of investigations, conducted in the Dept. of Mineral preparation at The Pennsylvania State University, which are directed toward the removal of ash and sulfur constituents from fine bituminous coal by size-density separations in cyclones, and toward a fuller understanding of the complex mechanisms involved in such separations. The purpose of this investigation has been to vary a series of dimensional parameters with an experimental 15-in. diam hydrocyclone, and to determine the relative effects of parameter variations upon ash and sulfur removal. Of special interest was the response to cone angle variation, since it is the high included cone angle that generally distinguishes the hydrocyclone as a gravity separator rather than a classifier. DESCRIPTION OF THE HYDROCYCLONE Fig. 1 is a dimensional drawing of the hydrocyclone used in the investigation. The unit was designed and fabricated by Heyl and Patterson, Inc., for experimental use. The dimensional parameters considered were: 1) Three cone sections having included angles of 45, 60, and 95,degrees, 2) Three 3 1/8 in. diam vortex finders having lengths of 15, 17, and 19 1/2 in., 3) Three 4 1/16 in. diam vortex finders having lengths of 15, 17, and 19 1/2 in. 4) Two apex sections having diameters of 2 and 2 1/4 in. The feed inlet orifice had a fixed diameter of 3 in. The hydrocyclone test installation was made at The Pennsylvania State University pilot plant at Barnes-boro, Pa., which operated at 50 tph. During the period that the hydrocyclone was tested, research was in progress at the Barnesboro site to reprocess refuse material from a near-by coal preparation plant in

Jan 1, 1968

By C. J. McHargue, J. C. Ogle

Lightly deformed columbiun single crystals which contained only parallel hoins or purullel and intersecting trains were annealed at 1000&apos; and 1600"C. No re-crystallizntion occurred in specimens hawing only parallel twins. Only noncoherent twin boundaries nzipated at 1000°C but both coherent and noncoherent ones moved al 1600°C. Recrystallization occurred within a few minutes at twin intersections at 1000°C. The orientation 01 the recrystallized grains differed front that of both the matrix and deformation twins, but could he derired by (110) and/or(112) rotations. ALTHOUGH twinning in metals has been extensively studied, there have been no definitive studies of the annealing behavior of crystals containing deformation twins. Some effects observed after annealing deformation twins have been summarized by Cahn1 and Hall2. Any or all of these phenomena are observed: 1) The twins may contract so that the sharp edges of the lens become blunted, and eventually the twin may disappear entirely. 2) The twins may balloon out at an edge, giving rise to a large grain having the same orientation as the twin. 3) The specimen may recrystallize; i.e., new grains are nucleated and grow at the expense of the twins and the crystal immediately adjoining the twin. Such grains have orientations which are not present before. Contraction has been observed in iron,3 titanium,3, 4 beryllium,5 zinc,8, 7 Fe-A1 alloy,&apos; and uranium.9 Long anneals at high temperatures are required to have any appreciable effect in these metals and only thin twins are absorbed. Lens-shaped twins are absorbed from the edges: the thin, almost parallel-sided twins are usually punctured in several places and each piece contracts independently. Absorption is very gradual and no sudden cooperative jumps have been observed. The expansion of a twin into a larger grain of identical orientation is unusual, but such growth has been observed in iron,"&apos;" zinc,6 and uranium." Crystals which have been deformed simultaneously by slip and twinning recrystallize first in the area adjacent to the twin. New grains appear faster where the twins intersect: but isolated twins, especially if thick, can also give rise to new grains. This type of recrystallization occurs in zinc.6, 7, 12, 13 and beryllium.14 Reed-Hill noted, in a single crystal of magnesium, the nucleation of a recrystallized grain at a twin intersection which had the same orientation as the second-order twin and which grew into the highly strained matrix.15 Short-time annealing has been reported to cause no change in the deformation twins in vanadium,16 columbium, 17, 18 tantalum,19 tungsten,&apos;&apos; and zinc.7 The purpose of this investigation was to note the effects of annealing on the coherent and noncoherent boundaries of deformation twins in columbium and to locate the nucleating sites for recrystallization. The orientation relationships, which the new recrystallized grains have with the parent crystal and the deformation twins, were also determined. EXPERIMENTAL PROCEDURE Single crystals of columbium were obtained by cutting large grains from electron-beam-melted buttons which contained 10 to 50 ppm C, 10 to 100 ppm O,, 1 to 10 ppm H2, and 10 to 15 ppm N2. The crystals were hand-ground and chemically polished until all grain boundaries were removed. The specimens were mounted in an epoxy resin and a face of each crystal was mechanically polished on a Syntron polisher using Linde A and then Linde B polishing compounds. After all faces were mechanically polished, the crystal was electrolytically polished to remove all distortion due to cutting and grinding. Laue photographs were taken of all faces of the crystals to determine the quality and orientation of each crystal. The crystals were compressed about 10 pct at -196 C in a specially constructed compression cage with an Instron tensile machine. Each crystal was separated from the top and bottom anvils by teflon films which acted as a lubricant. With the specimen crystal in position, the entire cage was cooled to -196°C by being submerged in a Dewar containing liquid nitrogen. The crystals were compressed at a rate of 0.02 in. per min and the load was recorded on a strip-chart recorder. After deformation the crystals were mechanically polished on 600-grit paper and Pellon cloth with Linde A and Linde B polishing compounds. The crystal faces were chemically polished and then etched. The twin planes were identified metallographically from an analysis of the twin traces on two surfaces. Annealing was carried out by placing each crystal in a columbium bucket made from the same electron-beam-melted material as the crystal itself and suspending the bucket by a tantalum wire in a quartz tube. After a vacuum of 10-7 Torr was attained, a furnace at 1000" or 1600 C was raised into position and the crystals held for various lengths of time. The crystals were repolished and etched after annealing to remove any surface contamination. Approximately 0.010 in. was removed during this process. The resulting surface was examined metallographically for microstructural changes due to annealing. A microbeam Laue camera mounted on a Hilger Micro-focus X-ray unit was used to determine the Orientstions of the recrystallized grains. This X-ray micro-beam camera had a 0.002-in.-diam collimator and incorporated the ideas of both and and chisWik21 and Cahn.22

Jan 1, 1967

By Stephen G. Dardaganian

Pore volume compresibilitieu measured in the laborutory on core .samples are reported for typical reservoir .sarrrl.rtorres at reservoir pressure. These cornpressibilities ore different for each reseriloir sample and cannot be correlated to porosity. The compressihilities are a1so functions of presslire. In recent years several authors have pointed out the importance of including pore volume compressibility of the reservoir rock in certain reservoir engineering calculations. Hall&apos; showed that in calculating the hydrocarbon volume of an undersaturated reservoir from the production per unit change in reservoir pressure, the neglect of rock compressibility could, in the extreme case of low porosity rock, lead to results in error by a factor of two. Hawkinsl&apos; and Hobson and Mrosovsky showed that both rock and interstitial water compressibilities must be included to achieve satisfactory accuracy in this calculation. Although rock compressibility is an important reservoir property in these calculations, very few data are available. Hall&apos; made compressibility measurements on a few samples in a small pressure range. No other direct pore volume compressibility measurements have been reported. Geertsma&apos; discussed the elasticity of reservoir rock but did not present any new data. In this paper, rock compressibilities suitable for reservoir calculations are reported for a variety of sandstones typical of U. S. reservoirs. The measurements were made through a range of pressures which is also typical of reservoirs. EXPERIMENTAL PROCEDURE Plugs 1 in. in diameter and 2- to 3-in. long were diamond drilled, 14 from cores from seven oil-bearing or potentially oil-bearing sandstone, one from a quarried sandstone, and another from a shallow, oil-free, pure orthoquartzite. The plugs were extracted in toluene, vacuum dried, and then weighed. The plugs were then saturated with kerosene and reweighed. This gave pore volume. They were then again extracted in toluene. vacuum dried, and jacketed in 0.004-in. thick copper foil. A short length of -in high pressure tubing led from the jacket. The jacketed sample was then saturated with kerosene and placed in the hydraulic pressure cell as shown in Fig. 1. Movement of the mercury slug showed the change in pore volume of the sample under a given pressure change. Temperature was about 72" F for all measurements. Before taking any measurements, the external pressure on the sample was raised to 10,000 psig and kept there for several hours. This formed the thin copper jacket around the plug. 7-he pressure was then removed and the sample left at least overnight, during which time it recovered its original pore volume. An external pressure of 12,000 psig was then applied and held constant. The change in pore volume was noted as the internal pressure was raised in 1,000 psig increments to 10,000 psig. The internal pressure was then reduced to zero, again noting pore volume changes,

Florida State Geological Survey, Tallahassee, Fla Herman Gunter, State Geologist A list of publications will be sent upon request Citizens of the state may obtain publications available without cost, others may have them for postage, or they can be sent express collect Twenty Annual Reports have been issued from 1908 up to and including the 21st-22nd Annual Report (1931). Only three are now available; 4th Annual Report (1912), 15th (1924) and 18th (1927) Reports since 1921 contain statistics on mineral production. All reports contain papers on various subjects, many of which are still available separately, and are so indicated in the list of publications. Seven bulletins have also been issued, principally of interest to paleontologists Bulletin 7, The Pensacola terrace and associated beaches and bars in Florida (1931), is of more general interest Bulletins will be sent upon receipt of postage. Thirteen Press Bulletins, most of which are available, have also been issued, but are of minor interest

Jan 1, 1933

By Thomas Leonard Watson

SULPHIDE ore-bodies of more or less lenticular shape occurring in metamorphic crystalline schists, gneisses, and. slates, and conforming closely in strike and usually in dip to the inclosing rock, have been recognized and described along the, Atlantic seaboard from Quebec to Alabama. The ore-bodies do not entirely conform, in all cases, to the structure of the inclosing schists, but they sometimes cut across the foliation in dip, though conforming more closely with the direction of strike. In composition the principal sulphide mineral is usually pyrite or pyrrhotite; or both, accompanied by copper, frequently gold, and almost always some sphalerite and galenite. Other. metallic ores are, so far as these sulphide bodies have been investigated, less common; silver is not uncommon, and in some of the pyrrhotite masses nickel is present in quantity sufficient to be commercially valuable. Cobalt frequently accompanies .the nickel, usually in smaller amount, and the two vary much in relative proportion. Quartz and calcite are probably the most common of the, non-metallic minerals. A number of silicate minerals are frequent representatives in one place or another; garnet, amphibole, biotite, pyroxene, zoisite and others. The sulphide bodies of Virginia, comprising both pyrite and pyrrhotite, are among the most extensive along the Atlantic seaboard. Among the pyrite bodies may be mentioned the well-known and extensively worked mines of Louisa and Prince William, counties, located on&apos; the main pyrite belt (A, Fig. 1) of the middle and northern, eastern Piedmont region in Virginia, and among the pyrrhotite bodies is the famous " Great Gossan Lead "(B, Fig. 1) of the Floyd-Carroll-Graysou coun-

Sep 1, 1907