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By N. Van Wingen, Norris Johnston

Laboratory flood pot testing of California sands has progressed to a considerable extent in the past 18 months. Flood evaluations have been carried out on over 200 large core samples. Many of these were heavy oil sands of high permeability and completely unconsolidated in nature. The oil frequently formed a bank, though some of the oil was recovered in the subordinate phase of the flood, by viscous drag. Flood pot recoveries as high as 1400 bbl/acre ft have been recorded. Reservoir analysis suggests a conformance factor of 0.4 to reduce laboratory recovery to probable field practice. Oils with viscosities up to 1800 cp have been successfully handled in flood pot evaluations. The shallow, loose sands are not well adapted to the application of- high pressures to offset the high viscosities. INTRODUCTION Secondary recovery may be said to have started 60 years ago when accidental floods occurred in the Bradford sand in Pennsylvania. About 1921 artificially conducted water drives came into extensive use and since that time the great Bradford field has been almost completely subjected to water flooding. During the last 30 years, most of the known medium and deeper production in California has been discovered and is being exploited by primary recovery methods supplemented in some instances by high pressure gas injection. The California area is just beginning to feel the need for secondary recovery in view of an unprecedented market demand and the rapidly rising cost of new pool discoveries. With the presently recognized desirability of secondary recovery in California, there must also be appreciated a number of serious differences between the water flooding problems here as compared to the territory east of the Rockies. California sands are generally thicker, and are frequently soft and argillaceous. The oils are often heavier and asphaltic. Much of the oil is below 15°API, occurs at shallow depth, is cool and free from appreciable dissolved gas, which results in relatively high reservoir oil viscosity. Secondary recovery is particularly beneficial where primary recovery has been poor and where no natural water drive exists. These conditions apply particularly to the heavy, shallow, clean production from soft, often argillaceous California sands so abundantly found at depths less than 1500 feet. Often, too, there is a totally insufficient supply of water of satisfactory quality to inject at a reasonable cost. Also, the crude oils are priced far below the premium Bradford crude. Although these and a number of minor problems beset the operator desirous of starting secondary recovery, great progress has been made in the past few years in finding how to adapt previous Mid-Continent and Eastern experience to water flooding in California. There are about nine projects for subsurface injection of water which can be said to classify as secondary recovery operations. Subsurface water disposal would so classify when the sand receiving the water is a nearby oil producer, as is often the case. When water is injected subsurface into a barren sand, the operation does not classify as secondary recovery. Several of the most active operators avail themselves extensively of preliminary engineering and laboratory work to guide their decisions, while others enter small scale flooding operations directly in the field. It is the laboratory work pertinent to several of the California secondary recovery projects that this paper discusses. PURPOSES OF LABORATORY FLOODING TESTS Experience in areas where water flood operations have been carried out has indicated that careful engineering planning is an important requisite for subsequent economically successful field operation. Floods that fail are more frequently those where operations were instigated without a prior engineering investigation to determine the effectiveness of the injection fluid as an oil displacing medium. Laboratory data are essential in the evaluation of an oil property for secondary recovery possibilities. Success or failure of secondary operations can under certain special circumstances be determined directly by cores and their subsequent routine analysis. This is particularly the case where flushing of the cores in the course of coring is negligible and where the results of the analysis can be compared with existing secondary recovery operations. Where these conditions cannot be' fulfilled, the application of core analysis is more limited. In such event, the results obtained by water flooding core samples in the laboratory have been found to be of prime importance. Cores may be flooded "raw" as taken from the well or in the event flushing and depletion of the cores in the process of drilling are major factors the fluid content may be artificially restored prior to the flooding. Laboratory studies should also be made to determine the suitability of the water selected for injection. Thus interaction between injected and formation water may cause precipitates to be formed which may plug the sand. Even more important, especially to California operations, is the possibility of the hydration of formation clays by the injection water. The aims of flood pot and associated tests are basically to determine the residual oil saturation after flood, the water-oil throughput ratio and to establish whether an oil bank is formed. Additional information which can be obtained from flood pot tests pertains to the pressure differential required to effect displacement, the relative permeability to oil in the oil bank and the relative permeability to water in the watered out region behind the bank.

Jan 1, 1953

By Charles Catlett

THE rapid development of the Portland-cement industry implies that the country is being very carefully searched for material suitable for its manufacture. Such material can be found at a great many places ; but deposits thoroughly satisfactory in chemical composition, and at the same time meeting all the other requirements connected with the economic production and profitable marketing of the product, are exceedingly few, if I may judge from the investigations which I have had to make in a number of States. The raw mixture which by its clinking and grinding forms Portland cement contains approximately 75 per cent. of carbonate of lime. This is found in nature as limestone or marl. The natural rock commonly carries either more than the required percentage of carbonate of lime (in which case it has to be reduced by suitable admixture of clayey ingredients) or less-in which case purer lime carbonate has to be added. The controlling ingredient, therefore, is carbonate of lime, or limestone of the right composition. Among the most serious difficulties is that of finding limestones low enough in magnesia to meet the exacting requirements of certain specifications. All engineers recognize that there is a limit beyond which it is not safe to let the magnesia go; but within that limit there is considerable difference of opinion. Until that difference is finally settled, no one would be willing to invest the large sum of money necessary for a Portland-cement plant, if it were not certain that the usual requirements in this particular could be satisfied. The usual limit in well-established specifications is 5 per cent. of magnesium oxide in the finished cement, but many specifications limit the amount to 3 per cent. These figures

Jan 9, 1907

By C. A. Bruch, W. E. McHugh, R. W. Hockenbury

EXPERIMENTAL results of the last decade have shown that both accelerator particles and reactor radiations produce significant changes in the properties of metals. These changes, called radiation damage, are important from both academic and engineering viewpoints. An understanding of the radiation damage produced by reactor radiations is difficult, because many kinds of particles having varying energies are present. Therefore, many fundamental studies are being conducted using accelerators in order to obtain nearly mono-energetic beams of various charged particles. Furthermore, many of these experiments are conducted at low temperatures in order to freeze in the radiation-produced lattice defects, after which the specimens are annealed to study the movement of the defects. It is found that although much of the low temperature radiation damage disappears by annealing at room temperature, a significant amount remains. Because there is no assurance that irradiation followed by annealing at a higher temperature will produce the same effect as irradiation at the higher temperature, and because there is an immediate interest in reactor radiation effects during exposures at and above room temperature, a major program is in progress at this laboratory to investigate some of these effects. This paper, representing a part of the program, compares the radiation damage to a group of metals which were exposed under nearly identical irradiation conditions. The metals, representing the three major crystal systems, include copper, nickel, titanium, zirconium, iron, molybdenum, and type 347 stainless steel. The effects of reactor radiations have been treated from a theoretical viewpoint by Seitz' and Ozeroff.' The problem considered by them has been the atomic displacements produced in an infinite medium by mono-energetic beams of both neutrons and charged particles. For reactor experiments, this treatment is not wide enough in scope, because specimens and reactor components are finite, and there are continuous energy spectra of the various particles. Therefore, the theoretical considerations will be extended in this paper to cover reactor conditions and the findings will be compared with the observed radiation damage. Experimental Materials—The six basic metals for this investigation were of commercial or higher purity, and the type 347 stainless steel was of commercial quality. These materials were supplied in various forms ranging from ingots to finished rods, and were fabricated to rods approximately % in. in diam and annealed for specimen stock. A summary of data pertinent to these materials is shown in Table I. The values for chemical compositions include both actual analytical values and typical values supplied by the manufacturer. More information concerning typical

Jan 1, 1957

By B. Ladanyi, B. Hoyaux

A first requirement for assessing the safety of a rock slope, either artificial or natural, particularly if the rock behaves in an elastic-brittle manner, is a propw knowledge of the distribution of stresses within the slope and their development with time and advancement of workings. There are a uumber of different methods that can be used, in principle, for obtaining this information, such as (1) stress measurement in situ, (2) photoelasticity,1 (3) analytical methods, and (4) numerical methods. In cases where the stress distribution cannot be obtained by classical analytical methods, as in the one considered herein, either photoelasticity or numerical methods are norrnally used. However, actually, due to the rapid development of high-speed computers, increasing preference is given to the numerical methods. Among the latter, the most successful ones, to date, for determining the stress distribution in slopes have been (1) the finite difference method,2-4 (2) the lumped parameter method,5 and (3) the finite element rnethod.6 However, for determining gravity stresses in a natural rock slope, only the finite difference method has been used until now.7,8 A solution of the same problem by the finite element method will be shown here in which a particular reference is given to the development of tensile fracture zones within the slope. FINITE ELEMENT METHOD The stress analysis using the finite element method originates from an extension of the matrix analysis of structures.9 It consists in decomposing a continuous medium, assumed herein to be homogeneous and isotropic,

Jan 1, 1972

By Reger, B. David

FOR many years, the belief that the so-called "Clinton" sand of Ohio might produce. oil and gas, if it could be reached, has been prevalent in West Virginia. To test this theory, many hundreds of thousands of dollars have been spent and several extremely deep wells, including the deepest boring of the world, have been sunk at various points. With the exception of one well now completed through the sand and the possible exception of another still in progress, all of these wells failed to reach the goal, either through failure of funds or through drilling accidents. REVIEW OF DEEP DRILLING The subject of deep-sand oil and gas in West Virginia has already been discussed by the author1 and by White.2 In these papers, it was shown that the Upper and. Middle Devonian shales, lying between the Berea oil sand (basal Mississippian) and the Oriskany sandstone (Lower Devonian), thicken steadily from eastern Ohio across northern West Virginia and Pennsylvania; but it was stated that the same rate of thickening did not occur in that part of West Virginia south of the Great Kanawha River and in the adjoining portion of Kentucky. In spite of this great eastward thickening, the fact that both the Oriskany and "Clinton" sands, the latter of which is equivalent to the White Medina of New York, exist in great thickness all along the Appalachian Mountains farther east, was held as proof of their probable presence in the deep Appalachian basin.

Jan 3, 1925

By Robert E. Delavault, Harry V. Warren

Circumstances determine whether it is better to make analyses in the field or in a permanent laboratory. The rubeanic acid test described in this article has been designed primarily for field use: it is simple and virtually foolproof, and it requires a minimum of field kit.* It is sensitive, easily detecting 4 ppm of readily extractable copper in a soil. This is by no means a quantitative test, but it is accurate enough to provide a valuable indicator of copper anomalous areas for both prospectors and field geologists. The easiest method for detecting metal deposits that do not produce visible float or stains is to make a simple chemical test for the metal in overlying soil, or in the silt of a stream that may have picked up metal farther upstream.

Jan 11, 1958

By John B. Gayle

PRESENT flotation processes depend almost entirely on the buoyant properties of air bubbles to effect separations of mineral and gangue, but there is no convenient method for measuring aeration in flotation cells. Consequently the relation between aeration and cell performance has not been definitely established. In connection with a comparative study of kerosene and froth flotation processes, a modification of the displacement method' was developed in which air escaping from the cell surface is measured with a wet-test meter. The arrangement shown in Fig. 1 is satisfactory for use with a number of cells and is not critical, except that care must be taken to use connections and tubing large enough to avoid excessive back pressures. Attempts to use smaller tubing resulted in back pressures that caused displacement of the liquid level in the receiver, so that some air was lost around the sides, an occurrence readily visible when measurements were taken with only water passing through the cell. With the arrangement shown, the manometer indicated an average reading of about ½ -in. water pressure. No loss of air was evident.

Jan 7, 1958

By Dennis V. D’Andrea, Sidney M. Runke

This paper describes the Bureau of Mines-El Paso Mining and Milling Company cooperative research program at the Emerald Isle mine3 near Kingman, Ariz. This research program was directed toward developing in situ leaching methods for 200,000 tons of ore exposed in the pit bottom and also 1,500,000 tons of ore under 200 feet of overburden adjacent to the pit. A test area in the pit bottom containing 15,000 tons of ore was blasted and leached. The results of this leach test were encouraging, and the entire pit bottom was leached in-place for about 7 months. Two test blasts under 200 feet of overburden and extending to 290 feet were detonated in an area near the open pit. The results of these blasts were evaluated by core drilling and permeability measurements. An in situ leaching system was designed, but not implemented, to recover copper from the 1,500,000-ton area near the pit. The company decided to terminate operations at the Emerald Isle mine in November 1975.

Jan 1, 1976

By William Fink

SINCE so many different curves have been published for the solid solubility of zinc in aluminum, it seems desirable to definitely establish the correct curve by two or more independent methods. The curve published in part I1 was based entirely upon X-ray measurements because the other methods employed at that time were not satisfactory. The present paper (the seventeenth of this series on high-purity aluminum alloys) describes the determination of the solid solubility of zinc in aluminum by the measurement of electrical resistivity at temperature and by microscopic examination. Since the publication of our part I in 1932, an investigation of the solubility relations in the aluminum-zinc system has been reported by Schmid and Wassermann2, who obtained their solubility results from lattice-parameter measurements on heat-treated and quenched speci-mens. The diagram that they published is shown in Fig. 1. In this investigation they also found by X-ray measurements that the parameter of the gamma phase changed continuously with zinc content and at such a rate that it would be identical with that of the beta phase at 60 atomic per cent zinc. It was concluded therefore that the beta and gamma phases, from the point of view of structural relationship, are one single phase. They also suggested a gap of immiscibility between 270° and 350° C. The complete boundary in this region could not be deter-mined satisfactorily by the method employed and therefore was not shown in their diagram.

Jan 1, 1936

By A. R. Troiano, E. A. Steigerwald, F. W. Schaller

The kinetics of crack propagation in a hydrogenated high-strength steel at subzero temperatures indicated that cracking progressed in a discontinuous fashion. The delayed failure process thus involves a series of crack initiations rather than the continuous growth of a single crack. The incubation time required for the initiation of delayed failure was controlled by the diffusion of hydrogen. The activation energy for the process was 9120 cal per g-alom. As the test temperature was lowered the incubation time constituted a greater percentage of the total tinze to fracture, and al -50°F the incubation and fracture times were essentially coincident resulting in instantaneous, catastroplzic fracture. A unique feature of hydrogen embrittlement is the delayed failure process in which the material fails under the action of a sustained load even though it is capable of supporting a higher load for a finite period of time.')' Delayed failure of steel and titanium is characterized by an incubation time, preceding the initiation of a crack in a highly localized region, and the subsequent growth of the crack to ultimate failure.3-5 The kinetics of crack growth in hydro-genated high-strength steels have been studied through the use of resistance measurements and have an important bearing on the mechanism of delayed failure. Delayed failure is believed to be merely a facet of hydrogen embrittlement in which sufficient time is allowed during a test for hydrogen diffusion. The current theories employed to explain hydrogen em-brittlement can be grouped into two types. one6-8 explains the embrittlement on the basis of the pressure exerted by molecular hdrogen in lattice voids or defects while the other9-10 attributes the embrittlement to the influence of hydrogen in the lattice. The mechanism used to explain the nature of the crack kinetics in delayed failure is based on the stress-induced diffusion of hydrogen to the area of maximum triaxiality.3 When the hydrogen concentration in this region reaches a critical value, a crack is nucleated. This initial crack propagates instantaneously until stopped by the higher fracture stress of the adjacent material of lower hydrogen content. The continuation of cracking must then await the further diffusion of hydrogen to the new region of triaxiality near the base of the crack where another initiation will occur. From this mechanism it follows that crack propagation is nothing more or less than a series of initiations with limited fracture. The purpose of this investigation was to examine the kinetics of crack growth in an effort to determine whether or not the process was discontinuous. Studies of crack propagation were conducted at low temperatures to retard the diffusion of hydrogen which is presumed to be the rate controlling process in delayed failure. MATERIALS AND PROCEDURE Specimen Preparation—A commercial heat of air-craft quality SAE-AISI 4340 steel, designated as Heat G, was furnished by the Republic Steel Co. in the form of %-in. hot-rolled bars. The material had the following chemical analysis: Composition of 4340 Steel—G Heat C Mn Si Ni Cr Mo 0,395pct 0.82pct 0.29pct 1.76pct 0.85pct 0.25pct Notched tensile specimens with less than a 0.001 in. radius, Fig. 1, were used throughout the investigation. The specimens were heat treated to a 230,000 psi strength level according to the following sequence: a) Normalize bar stock at 1650°F for 1 hr, b) Stress relieve at 1200°F for 4 hr and furnace cool,

Jan 1, 1960

By Frederick H. Williams

THE object of this paper is to bring to the attention of those members of the Institute who are interested in the subject, a method for the estimation of manganese, which has been frequently used in the laboratory of the Vulcan Steel Works during the last eight months, and which is deemed worthy of more general use. It is simply an adaptation of well-known processes, for the purpose of determining this element in pig iron and steel. In general, the method consists in precipitating the manganese as binoxide, dissolving in dilute sulphuric acid in the presence of an excess of a standardized solution of oxalic acid, and titrating the remainder of the oxalic acid left unoxidized with permanganate of potash. For this method, therefore,. two standardized solutions are needed, viz., a permanganate solution and an oxalic acid solution. The permanganate solution is made so that 1 c.c. is equivalent to 1 mgr. of iron. By using so weak a solution and keeping the bulk of the solution to be titrated down to 75 c.c. or a little less, the accuracy of the process is greatly increased, since the final reaction is almost if not quite as sharp as if a stronger solution were used. The oxalic acid' may be of almost any strength, but if made so that 1 c.c. requires 3 c.c. of the permanganate to oxidize it, it will auswer well. The process is then, as follows : from 1 to 2 grams of pig iron or steel are dissolved in 40 to 50 c.c. of concentrated nitric acid with the aid of heat. As soon as solution is effected, in the case of steel, add gradually and cautiously, the solution being kept boiling, sufficient pulverized chlorate of potash to precipitate the manganese as binoxide. Two grams will be a great abundance. In the me of pig iron, first remove the carbon and silica by filtering the solution through asbestos and washing with concentrated nitric acid. Filter the strong acid solution in which the binoxide has been precipitated, through asbestos with the aid. of the filter pump. When it has entirely passed through wash well with water. Blow the contents: of the funnel into the beaker in which the precipitation was made and rinse the funnel with a little water.

Jan 1, 1882

By T. Wasson

Midland and Isabella counties have been the centers of activities for the year 1932. The major part of the 6,967,000 bbl. produced during the year was obtained from the old Mt. Pleasant field and the Eastern Extension field. The total production to the end of 1932 for Michigan, from all fields since discovery, is 20,679,600 bbl. Production in the Eastern Extension field was obtained from Dundee limestone of Devonian age, at a depth of about 3600 ft. As a result of the year's activity in this area, the limits of the Eastern Extension have been defined. No new fields have been discovered. A number of prospective areas have been drilled and condemned by dry holes. Probably the most outstanding development of the year was the discovery of oil on the Yost farm, in sec. 35 of Greendale Township, Midland County, by the Pure Oil Co., in the early part of December. Production is from the Dundee limestone similar to that in the Eastern Extension but is 4 miles south of that area. The first well has averaged about 500 bbl. per day since completion. It is near the center of a solid block held by the Pure Oil Co., so that rapid development due to competitive drilling is not expected. During the early part of the year the Midland-Isabella area was under proration, but on May 17 proration restrictions were discontinued and remained so until the end of the year. While the Eastern Extension saw the principal activity, some work was done in other fields of Midland and Isabella counties, but very little new production was added. The Vernon field on the north showed a slight increase in production over 1931. The Leaton field decreased. The well in Porter Township, Midland County, is still kept in the production ranks but very little drilling was done in that area. The old Saginaw field, Saginaw County, still remains the only spot in the state where commercial production is obtained from the Berea sand. It showed a slight decline over 1931 with no new producers for the year. The Muskegon field on the west side of the state showed a considerable decline over 1931, but owing to the rejuvenation of some wells by the use of acid this drop was not as great as might have been expected. Only two new producers were drilled in that area.

Jan 1, 1933

By H. R. Tunnell

Every one connected with a mine knows that it is hard to keep down the costs of moving ore from the place where it is broken to the shaft or portal. Considered broadly, the subject of handling would cover all work done in a mine, but here it has been limited to handling ore in stopes and drifts, through chutes, and on levels. Good ventilation and a strict enforcement of rules for the prevention of dust safeguard the health of the men and lower transportation costs. Handling ORe in Stopes and Drifts The object of all stoping is to mine ore of good grade at low cost. In Butte, most of the ore is mined, from veins dipping from 60° to the horizontal to those which are vertical, by means of square-set, rill, timbered-rill, and back-filling stopes. When blasted, a part of the rock is carried by gravity to the grizzlies over the chutes, the rest is shoveled. Rock, when broken in a stope, may be deflected by inclined floors or slides toward the top of a chute leading to the level below. Ore broken clean enough to ship should be carried, by the force of its fall, into a chute or.as nearly into that chute as the safety of the stope and the men permits. Ore that requires sorting, about 80 per cent. of all ore mined, is caught on a grizzly and the waste picked out. Waste rock broken in a stope is used to fill that stope and should be blasted into the gob with as little handling as possible. About two-thirds of all the ore mined in Butte comes from square-set stopes; this method was used in nearly all cases until 1912, when the rill method of stoping was introduced in the Tramway mine, of the Anaconda Copper Mining Co. The shoveling and timbering costs in square-set stopes are high when compared with the costs in the rill stopes. All the ore cannot be carried to the chute, in falling from place, unless the chutes are very close together or the distance from the back of the stope to the top of the chute is sufficient to slide all the ore broken. The number of chutes is limited by the tendency of the ground to cave; if carried too close together the stope is weakened. Small stopes, or blocks,

Jan 1, 1923

By H. R. Tunnell

Every one connected with a mine knows that it is hard to keep down the costs of moving ore from the place where it is broken to the shaft or portal. Considered broadly, the subject of handling would cover all work done in a mine, but here it has been limited to handling ore in stopes and drifts, through chutes, and on levels. Good ventilation and a strict enforcement of rules for the prevention of dust safeguard the health of the men and lower transportation costs. Handling ORe in Stopes and Drifts The object of all stoping is to mine ore of good grade at low cost. In Butte, most of the ore is mined, from veins dipping from 60° to the horizontal to those which are vertical, by means of square-set, rill, timbered-rill, and back-filling stopes. When blasted, a part of the rock is carried by gravity to the grizzlies over the chutes, the rest is shoveled. Rock, when broken in a stope, may be deflected by inclined floors or slides toward the top of a chute leading to the level below. Ore broken clean enough to ship should be carried, by the force of its fall, into a chute or.as nearly into that chute as the safety of the stope and the men permits. Ore that requires sorting, about 80 per cent. of all ore mined, is caught on a grizzly and the waste picked out. Waste rock broken in a stope is used to fill that stope and should be blasted into the gob with as little handling as possible. About two-thirds of all the ore mined in Butte comes from square-set stopes; this method was used in nearly all cases until 1912, when the rill method of stoping was introduced in the Tramway mine, of the Anaconda Copper Mining Co. The shoveling and timbering costs in square-set stopes are high when compared with the costs in the rill stopes. All the ore cannot be carried to the chute, in falling from place, unless the chutes are very close together or the distance from the back of the stope to the top of the chute is sufficient to slide all the ore broken. The number of chutes is limited by the tendency of the ground to cave; if carried too close together the stope is weakened. Small stopes, or blocks,

Jan 1, 1923

By Thomas Clagett

THE Pocahontas coal field comprises the area in Tazewell County, Va., and Mercer and McDowell Counties, W. Va., in which Nos. 3 and 4 Pocahontas seams of bituminous coal are mined. It is a mountainous region with ridges rising 300 to 1000 ft. (91 to 305m.) above the very narrow valleys and, except for some poorly kept farms along the tops of the ridges and the area occupied by the mining and coking plants and villages, is wooded throughout. First shipments were made from Pocahontas in 1883. There are now, about 100 mines, 95 per cent. of which are drift mines. At the end of 1920, approximately 298,000,000 gross tons of coal had been produced and 35,000 acres of coal mined. No. 3 seam is from 4 to 10 ft. (1.2 to 3 m.) thick, but the change is gradual and the thickness fairly uniform in any particular locality. It has a streak of bone, about 2 in. (5 cm.) thick, io which the coal adheres on both sides. In parts of the field this bone is replaced by hard slate, usually from 2 to 4 in. thick. The seam has a slate or fireclay bottom and a draw-slate top. No. 4 seam is from 65 to 80 ft. (19.8 to 24 m.) above No. 3 seam, and of minable section only in the western part of the field. It is from 3 to 8 ft. thick and has two streaks off bone, each from 1 to 2 in. thick. It has a fireclay or slate bottom and is overlaid by from 4 to 24 in. of a laminated coal and slate, locally termed "black rash," which is separated from the coal seam by about 3 in. of slate. This "black rash" averages about 25 per cent. ash, but contains some clean coal. The coal in both seams is quite soft. Cleavage sufficiently well defined to exert an influence on the mining is rare. Overlying strata are slates, shales, and sandstones, the sandstones predominating, extending from .30 to 1000 ft. above No. 3 seam. The strata directly above the. draw slate of No. 3 seam, or "black rash" of No., 4 seam, and up to the first sandstone are usually slates or shales, about 10 ft. thick over the

Jan 5, 1922

By M. G. Corson

INFORMATION regarding the use of manganese alloys has hitherto been incomplete and available only from widely scattered sources. This paper attempts a systematic description of properties .and uses of alloys other than manganese bronze, manganin and duralumin, which are probably the only non-ferrous alloys associated with manganese in the minds of most metallurgists.1 PROPERTIES OF MANGANESE Manganese can react either acid or basic; it is hard to obtain in the pure state; it has two different crystalline lattice structures, both of the cubic system, one stable above 850° C. and the other below 650° C., both coexisting in the intermediate range. Although the commercial product is hard and brittle, the pure metal seems to be ductile. In fact Zhemchuzhny, in Russia, was able to obtain ductile manganese and draw it to fine wires by the simple expedient of adding about 3 per cent. copper to the alumino-thermic product. Copper most probably brings the silicides, aluminides, nitrides and carbides present in the commercial product into a much less harmful form of distribution. One industrial use that can be foreseen for wrought manganese articles is as anodes for electrolytic processes in sulfate and nitrate solutions, particularly with high voltages. A practically insoluble skin of manganese peroxide is likely to form in such cases and an insoluble and strong anode material will be available.

Jan 1, 1927

By J. L. Bray

ACCELERATED programs, as discussed in this paper, refer to the year-around operation of a college or university with three sixteen-week or four twelve-week terms per year, with pauses between sufficient only to allow the administrative group to perform their functions. In most engineering schools, under such conditions, the conventional four-year course can be condensed to two years and eight months. These accelerated programs of study, which now exist in most of our technical institutions, were forced upon us under the influence of the outburst of wartime patriotism, perhaps one might go so far as to call it hysteria. Many of us felt and hoped that these programs were temporary expedients, but the cold hard threat of permanency now rears its head; consequently we should examine the merits and shortcomings of the new order, and evaluate its effect on our educational system, since it may persist in some form in the postwar years, because of the changing political and financial situation. Many engineering students are being educated in state supported institutions

Jan 1, 1944

By N. A. Gokcen, J. Chipman

Aluminum and oxygen dissolved in liquid iron were brought into equilibrium with pure alumina crucibles and atmospheres of known H2O and H2 contents to study the reactions: 1—Al2O3(s) = 2 Al + 3 0; 2—Al2o3(s) + 3H2(g) = 2Al+ 3H2o(g); and 3—H2(9) +O = H2O(g). Aluminum strongly reduces the activity coefficient of oxygen and similarly oxygen reduces that of aluminum. Values of the product [% All" • [% O]3 are much smaller than those found in previous experimental studies and are of the order of magnitude of the calculated values. ALUMINUM is the strongest deoxidizer commonly A used in steelmaking, but the extent to which it removes dissolved oxygen has been debatable. The relationship between aluminum and oxygen has not been determined reliably not only on account of the usual experimental difficulties at high temperatures but also because of uncertainties in the analyses of very small concentrations of oxygen and aluminum. The earliest experimental attempt of Herty and coworkers' was followed by a more systematic study of Wentrup and Hieber.' These authors added aluminum to liquid iron of high oxygen content in an induction furnace and considered that 10 min was sufficient to remove the deoxidation products from the melt. Parts of the melts thus obtained were poured into a copper mold and analyzed for total aluminum and oxygen (soluble plus insoluble forms), assuming that the insoluble parts were in solution at the temperatures from which samples were taken. It is conceivable that the furnace atmosphere in their experiments, consisting of mainly air at 20 mm Hg pressure, was a serious source of continuous oxidation and therefore that their oxygen concentrations were correspondingly high. Scattering of their data was explained to be well within the maximum inaccuracy of 10°C in the temperature measurements and errors of ±0.002 pct each in the oxygen and total aluminum analyses. Maximum and minimum deoxidation values, i.e., values of the product [% All' . [% O] differed by factors of 10 to 15; mean values of 9x10-11 and 7.5x10-9 ere reported at 1600" and 1700°C, respectively. Hilty and Craftsv determined the solubility of oxygen in liquid iron containing aluminum, using a rotating induction furnace. Pure alumina crucibles used in their experiments contained the liquid iron which in turn acted as a container for slags of varying compositions consisting mainly of Al2O3, Fe2O3, and FeO. The furnace was continuously flushed with argon, and additions of aluminum and Fe2O3 were made in the course of each experimental heat. The inner surfaces of their alumina crucibles were covered with a substance other than pure Al2O3, containing both iron oxide and alumina. Although frequent slag additions can change the composition of slag in the liquid iron cup formed by rotation, the inner surface of the crucible must depend upon the transfer of oxygen or aluminum through the liquid iron for any adjustment in composition. It is not clear that their metal was in equilibrium with the crucible wall, but it is clear that it was not in equilibrium with Al2O3. Their deoxidation product, [% A].]" • [% O]3, varied by a factor of more than 50; the average values of 2.8x10- and 1.0x10-7 were selected for temperatures of 1600" and 1700°C, respectively. Aside from the experimental determinations, attempts have been made to calculate the deoxidation constant for aluminum indirectly from thermody-namic data. Schenck4 combined the thermodynamic data for Al2O3 and dissolved oxygen in liquid iron by assuming an ideal solution. His calculated values are 2.0x10-15 and 3.2x10-13 at 1600" and 1700°C, respectively. Later, Chipman5 attempted to correct for the deviation from ideality and derived an expression which led to deoxidation values of 2.0x10-14 and 1.1x10-12 at 1600" and 1700°C, respectively. The errors in these treatments originate mainly from inaccuracies of thermal data and uncertainties regarding the activity coefficients of dissolved oxygen and aluminum. The purpose of this investigation was to study the equilibria represented in the following reactions in the presence of pure alumina: Al2O3(s) = 2Al + 3O K = aAl2.ao3 [1] Al2O3(s) + 3H2(g) = 2Al + 3H2O(g) H2O K2 = aAl2(H2O/H2 ) [2] H2(g) +O = H2O(g) K3 = 1/ao (H2) [3] The experimental method consisted of melting pure electrolytic iron, usually with an initial charge of aluminum, in pure dense alumina crucibles under a controlled atmosphere of H,O and H2 and holding

Jan 1, 1954

By W. W. Whitaker, C. W. Manry, R. H. McLean

In an eccentric annulus, cement may favor the widest side and bypass slower-moving mud in the narrowest side. Tendency of the cement to bypass mud is a function of the geometry of the annulus, the density and flow properties of the mud and cement and the rate of flow. Bypassing can be prevented if the pressure gradient protluced from circulation of the cement and buoyant forces exceeds the pressure gradient necessary to drive the mud through the narrowest side of the annulus at the same velocity as the cement. In the absence of buoyant forces, one requirement for this balance is maintenance of the yield strength of the cement greater than the yield strength of the mud multiplied by the maximum distance from the casing to the wall of the borehole and divided by the minimum distance. If the yield strength of the cement is below this value, bypassing of mud cannot be prevented unless buoyant forces or motion of the casing significantly aid the displacement. INTRODUCTION Successful primary cementing leaves no continuous channels of mud capable of flow during well treatment and production. Prevention of channels requires care. Tep-litz and Hassebroek provide evidence of channels of mud after primary cementing in the field.' Channeling of cement through mud in laboratory experiments has also been reported.'-' Recommendations for improving the displacement of mud include (1) centralizing the casing in the borehole,'-" 2) attaching centralizers and scratchers to the casing and moving it during displacement,18 "3) thinning the isolating the cement by plugs while it is circulated down the casing,%( (5 establishing turbulence in the cement," and (6) holding the cement slurry at least 2 lb/gal heavier than the mud and circulating the cement slurry at a very low rate of flow.' Although much has been written about the above parameters, the relative importance of each has not been well defined. In this investigation, the mechanics of mud displacement are described through results from analytical models and experiments. The model chosen — a single string of casing eccentric in a round, smooth-walled, impermeable borehole — is analagous to casing centralized in a borehole which is not round and to placing more than one string of casing in a borehole. In each, some paths for flow are more restricted than others. A fluid flowing in the borehole may seek the least restricted, or most open, path. This tendency for uneven flow can lead to channeling of cement through mud unless preventive measures are taken. The analytical models describe channeling and give means of balancing the flow. Experimental data test the analytical models and illustrate effects of motion of the casing, differences in density and mud's tendency to gel. Results are encouraging. Piston-like displacement of mud by an equal density cement slurry is possible through proper balance of the flow properties of the mud and cement slurries to the eccentricity of the annulus. The more eccentric the annulus, the thicker must be the cement relative to the mud. If proper balance is not achieved. bypassing of mud by cement cannot be prevented without assistance from motion of the casing or buoyant forces. Increasing the rate of flow can help to start all mud flowing but cannot prevent channeling of cement through slower moving mud in an eccentric annulus. Thinning the cement slurry tends to increase channeling although the extent of turbulence in the annulus may be increased. Description of flow in an eccentric annulus begins in the next section. It is assumed that (1) the casing is eccentric and is stationary, (2) the mud and cement slurries have the same density and (3) the gel structure of the mud has been broken and the mud and cement follow the Bingham flow model. Effects related to these restrictions will be discussed. FLOW PATTERNS SlNGLE FLUID IN ANNULUS Flow of a single fluid through an eccentric annulus is illustrated in Fig. 1. Part A shows laminar flow of a Newtonian fluid. This distribution of flow was calculated by Piercy, Hooper and Winney.' In fully developed turbulent flow, the velocity distribution around the annulus is less distorted, but the flow still favors the widest part of the annulus Parts B, C and D of Fig. 1 are a qualitative representation of the flow of a Bingham fluid. The yield strength of the fluid increases the severity of bypassing compared to Newtonian flow. At a very low rate of flow, all flow is confined to that portion of the annulus which has the minimum perimeter-to-area ratio. The fluid shears on the perimeter of that area when the pressure gradient multiplied by the area just exceeds the yield stress of the fluid multiplied by the perimeter. Whether or not the minimum perimeter-to-area region encompasses all of the annulus or only a part (as shown in Part B) depends on the geometry of the annulus. If only a part begins to flow, increasing the rate of flow increases the area flowing until finally there is flow throughout the annulus.

By J. H. Slater

There is nothing particularly new about the use of sinter in a blast-furnace burden. For many years flue dust has been sintered at the various blast-furnace plants to put it in a form that could be recharged to the furnace with some assurance that it would stay in the second time. This paper is not a compilation of scientific data but only an attempt to give our experience with sinter. There are some unexplained discrepancies in the tests given here, such as the flue dust produced and iron tonnages, but the results given are the results that were obtained. Sinter is not the cure-all for blast-furnace problems. A good coke is still the most important factor in good furnace operation. However, good sinter is a big help. As flue dust is produced under varying operating conditions and fluctuates greatly in carbon content, it is difficult to produce from flue dust alone a uniform sinter, which will give satisfactory blast-furnace results. Sinter made from flue dust alone will change in character with every carbon variation. Almost invariably flue dust contains carbon in excess of the amount necessary to make good sinter, and this excess must be burned out. This, of course, necessitates reduction of the speed of the sintering machine and thus a decrease in tonnage. An alternative is to dilute with a carbon-free material to use up the excess carbon. High-carbon material when sintered results in a very dense, undesirable product, since the excess heat melts the charge into a hard, nonporous mass, and tends to form refractory silicates. At the Republic Steel plant in Cleveland, there had existed since 1924 a continuous-type sintering machine having 224 sq. in. of grate-bar area, capable of producing 9 to 11 tons of sinter per hour from flue dust. Under a former management, it was the practice to sinter only the flue dust from the blast furnaces, no attempt being made to recover the gas-washer sludge. As the dust was limited, the plant operated only part time. In other words, there was a sintering plant representing considerable investment, with good facilities for screening and preparation of raw materials, and flue dust available for only part-time operation. The sinter produced was costly and not a very desirable product. The first step was to install equipment for the recovery of the gas-washer sludge. This was done in the form of two large thickeners and a filter, which have proved very satisfactory. It was then decided to screen out the minus 3/8-in. fines from the Mesabi ores, mix them with flue dust and coke breeze when necessary, and operate the plant at full capacity. The tonnage increased to 18 tons per hour and the product was much improved. Sintering costs fell and the product was a preferred blast-furnace material, which on an iron unit cost compared very favorably with good Mesabi ores. On the average, 40 per cent flue dust and 60 per cent fine ore was being used. The oversize from the ore screening also became a preferred product for the

Jan 1, 1941