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By Jack D. Webber

The successful use of oil well packers requires, in part: an understanding of the pressures which exist at the packer in various applications and an understanding of the characteristics of the various types of packers. It is with these pressures, the resultant forces, and the characteristics of packers. that this paper is primarily concerned. An oil well packer may be defined as a mechanical device for blocking the passage of fluids in an annular space. In the more usual case, the annular space is that between the tubing or drill pipe in a well and the casing, and packers which block such an annular space are broadly referred to as casing packers. In the other case. the annular space is that between the tubing or drill pipe and the walls of an open hole, and packers for blocking this space are generally called formation packers. While the hydraulics involved are essentially the same for casing and formation packers. a greater variety of conditions are encountered in the use of casing packers and only casing packers will be discussed. After a packer has been set and a pressure seal effected between tubing and casing, the packer is comparable to a piston in a cylinder. Pressures acting upon a piston result in forces which will move the piston unless some means is provided to prevent such movement. In the same manner, pressures acting upon a packer will move the packer unless there is present a sufficiently great restraining force. PACKER CLASSIFICATIONS Packers may be classified according to the pressure conditions under which they are capable of blocking the annular space between tubing and casing. Fig. 1 shows schematically two types of packers in common use. These packers are capable of blocking the annular space against the passage of fluids under a differential pressure of significant magnitude only when the pressure in the annular space above the packing element is greater than the pressure below. It may be seen that in Fig. l-a. slips with teeth which bite into the casing and prevent downward movement are provided. In Fig. 1-h. an anchor prevents downward movement. In each case, there i-only the tubing to prevent upward movement when differential pressures act to move the packers upwardly. Packers which hold only a significant differential pressure acting downwardly have been in use since the early days of the oil industry and will hereafter be referred to as conventional type packers. In many packer applications operating conditions will 1.crult in differential pressures across the packer which will at times act to move the packer upwardly, and at other times, act to move the packer downwardly. For these applications, designs are available which will block the annular space and resist movement in either direction. Fig. 2-a shows schematically a packer of this type which is designed to be run into a well and set, and removed when desired by merely pulling the tubing. It will be noted that two sets of slips are provided-— one set above the packing element to prevent upward movement, and another set below the packing element to prevent downward movement. This packer is built around a mandrel which is essentially a part of the tubing. and which is free to move longitudinally within certain limits through the set packer. Fig. 2-b shows schematically a permanent type packer which is capable of holding pressures from either direction. Here again, two sets of slips are provided to prevenl movement of the packer. This packer is designed to become virtuallv a part of the casing when set and it is made of drillable material so that it may be drilled out when its removal is desired. The seal nipple shown effects a pressure seal between the tubing and the packer. This seal nipple is a part of the the tubing, and the nipple and tubing may be withdrawn from the well without disturbing the packer. It should be noted that these figures are not representative of all available packers which are designed to hold pressures from both above and below. Packers which resist movement in either direction will hereafter be referred to as universal type packers. There is a third type of packer in general use and this type is designed to block the passage of fluids when the pressure below the packing element ii greater than that above. This type is provided with slips which prevent upward movement of the packer and is somewhat similar to a conventional type packer run upside-down. Packers designed to hold pressure only from below are made in a variety of designs and are usually owned and operated by service companies.

Jan 1, 1949

By Walter Wood

The specifications for cast-iron pipe that have been submitted at this meeting are practically the outgrowth of those which were originally adopted, about 1860, by Mr. Kirkwood of Brooklyn, N. Y. They have been changed but little since that time, the mode of testing pipe in the hydraulic press, as well as the method of coating, not having departed from the original practice. The chief addition to the specifications has been in the adoption of a test-bar 2 in. by 1 in. in crosa-section, broken between 24-in. supports by a transverse strain applied on the flat side. Test-bars were first introduced into pipe-specifications about 1870. At that time a satisfactory test-piece had to stand a tensile strain of 16,000 lb. per square inch. The present test-bar was adopted in order to give some idea of the deflection of the iron that was being used. These deflection-tests have been so satisfactory to both engineers and manufacturers, and have met the requirements of the pipe-trade so accurately, that it is considered best to hold to the . 2-in. by 1-in. bar instead of adopting the short " Arbitration Bar," which yields but little information from the stand-point of deflection. The new feature in these specifications is the effort to reach standards to which pipe shall be cast in the future, thus permitting the manufacturer to make up stock and the purchaser to find material on hand when he asks for deliveries. Heretofore there has never been ally standard weight for pipes? although in some sizes certain weights are more frequently asked for than others. The fixing of standards for pipes permits the adoption of definite standards for fittings. It has been aimed to secure such uniformity for fittings as will permit the same class of fittings to cover two or more classes of pipe. In order to accomplish

Jan 1, 1905

By W. R. Hibbard

T has been proposed1 on the basis of slip and flow I- that the ideal deformation texture of cold rolled face-centered cubic metals is (110) [112]. As pointed out recently by Brick,' this theory does not account for the difference between the pole figures for cold rolled copper and 70-30 alpha brass, which develop strikingly different spreads from the ideal orientations. It does suggest that the pole figures might be similar for these two metals if they are rolled sufficiently. This has been found to be the case for cold drawn wire.' Experimental Procedure: Four inch sections of cast cakes of copper and 70-30 alpha brass were obtained through the courtesy of the Bridgeport Brass Co. Pieces from these cakes were cold straight rolled 99.993 pct reduction in thickness with the assistance of the Remington Arms Co., Bridgeport, Conn., and by the use of pack rolling. Pole figures were made using the X ray camera and techniques described by Brick.' These pole figures were not significantly different from those previously published for the same materials rolled 98 pct. Eight thicknesses of this strip were then pack rolled. This material failed completely and extensively in tension as shown in fig. 1, developing wavy transverse fractures and bulges. No further work could be done on these samples. Summary: Within the range of this experiment, no attainable amount of cold rolling would cause the pole figures of cold rolled copper and 70-30 brass to become similar, in spite of the fact that such a similarity is predicted by theory.' References W. R. Hibbard. Jr. and M. K. Yen: Trans. AIME (1948) 175, 126. Met. Tech. Feb. 1948. TP 2334. ²R. M. Brick: Trans. ASM (1949) 41A, 99. 3W. R. Hibbard, Jr.: Trans. AIME (1949) 185, 598, Jnl. of Metals, Sept. 1949. "R. M. Brick: Trans. AIME (1940) 137, 193. 'R. M. Brick, D. L. Martin and R. P. Angier: Trans. ASM (1943) 31, 675.

Jan 1, 1951

By G. S. Farnham, D. A. Scott

Partition of certain elements, particularly nickel, was determined for slowly cooled steels, the greater number containing from 0.30 to 0.35 pct C. Approximately 3 pct of the nickel, 18 pct of the manganese, 33 pct of the molybdenum, and 35 pct of the chromium occur in the carbide phase. Partition of one element is not much affected by the presence of another. THE object of this investigation was: 1—to determine the partition of nickel between cemen-tite and ferrite in a medium and high carbon series of annealed steels; 2—to evaluate the effect of chromium and/or molybdenum or both on the partition of nickel in medium carbon annealed steels; 3— to determine the partition of chromium between ferrite and carbide in a series of medium carbon annealed steels; and 4—to study the chemistry of the carbides formed in an AISI 2330 steel during both isothermal transformation, and quench and draw experiments at 1150°F (620°C). The partition of alloying elements between ce-mentite and ferrite in annealed steels has been the subject of many investigations, a summary of which has been given by Austin1. However, there is no satisfactory data available concerning the partition of nickel. It has generally been assumed that virtu- ally all of the nickel concentrates in the ferrite phase in annealed steels. Two investigators"." have reported the finding of nickel in extracted carbide residues. Recently, it has been shown4 that nickel substitutes up to 40 pct for iron in the lattice of cementite produced by the low temperature carburization of iron-nickel powders. In addition to the lack of data on nickel partition, the literature also fails to make reference to the effect of other elements on the nickel content of the carbide in nickel-bearing steels. The partition of molybdenum between ferrite and carbide has been reported"-' for hypoeutectoid steels, isothermally transformed at 1300°F (705°C) and

Jan 1, 1954

By E. S. Bardwell

Each of the three refining furnaces in use at Great Falls is provided with a waste-heat boiler. The general arrangement of furnace and boiler is as shown in Fig. 1. Two of the furnaces have hearths 45 ft. 7 in. by 14 ft. 6 in.; the third has a hearth 46 ft. 101/2 in. long and of the same width as the other two. The furnaces at present are equipped for firing with natural gas. As originally constructed two of the furnaces had hearths 36 by 14 ft., and were designed to use coal on grates, but these furnaces were lengthened and fired successively with pulverized coal and oil. The third furnace was added in 1926 and was oil-fired until 1928. Since that year natural gas has been in use on all of the furnaces. The changes that have occurred in connection with furnace design and the comparative performance of the furnaces with various fuels were described1 by the author in 1932. Two of the furnaces are provided with Stirling class A-30 boiler. rated at 400 hp. and having 4008 sq. ft. of heating surface. The thir furnace has a class A-21 Stirling waste-heat boiler rated at 300 hp. with 2805 sq. ft. of heating surface. The boilers are not baffled. The waste-heat boilers are placed on a platform 17 ft. 3 in. above the level of the furnace charge floor. The waste gases from the furnace pass by a vertical uptake flue, 4 by 3 ft. inside at the point where it enters the chamber, to a brick chamber from which the gases pass over a bridge wall and into the boiler setting. The uptake flue and chamber are constructed of silica brick. Each furnace has a steel stack 125 ft. high, measured from the boiler floor. The stacks have a 41/2-in. firebrick lining and are 5 ft. in diameter

Jan 1, 1934

By William Carton, Cyril U. Cooledge

BecauSe the rich La Colorada orebody of the Cananea Consolidated Copper Co. does not outcrop at surface, after its discovery (by churn drill) and before mining was begun, a large amount of development work was done to learn the shape and extent of the ore and to provide for economic and effective mining. As these openings progressed, the actual shape of the orebody became known and with this knowledge came the realization that mining this rich and unusual deposit without loss of ore or danger to workmen presented several difficult problems. This paper will describe one of these problems; namely, the handling of an inverted cone-shaped core of barren rock, on the periphery of and below which the ore occurred. It mas clear that as the ore was mined out from around and under this inverted cone of rock, weighing about 3,000,000 tons, it would hang over the workings, held suspended only by its own tensile strength and threatening to crash into the stapes below with disastrous results to men and mine. The method of mining developed during the years 1928,1929 and 1930 anticipated and provided for the eventual collapse of this core of rock. In November 1936 the core settled down gradually on the edges of the stopes, descending a total of about 50 ft. vertically. The men in the stopes had ample warning and no lives were endangered. There were no air blasts in the mine. The core broke loose from its top support at the 500-ft. level along old fracture planes and evidently is crushed to the ninth level. Movement was of short duration; the following shift went to work in the stopes as usual. No ore was lost. Geological Structure The shape of the orebody is illustrated in Fig. 1. The geology of the Colorada orebody has been described by V. D. Perry1 and by W. G.

Jan 1, 1940

By Cyril U. Cooledge, William Carton

BecauSe the rich La Colorada orebody of the Cananea Consolidated Copper Co. does not outcrop at surface, after its discovery (by churn drill) and before mining was begun, a large amount of development work was done to learn the shape and extent of the ore and to provide for economic and effective mining. As these openings progressed, the actual shape of the orebody became known and with this knowledge came the realization that mining this rich and unusual deposit without loss of ore or danger to workmen presented several difficult problems. This paper will describe one of these problems; namely, the handling of an inverted cone-shaped core of barren rock, on the periphery of and below which the ore occurred. It mas clear that as the ore was mined out from around and under this inverted cone of rock, weighing about 3,000,000 tons, it would hang over the workings, held suspended only by its own tensile strength and threatening to crash into the stapes below with disastrous results to men and mine. The method of mining developed during the years 1928,1929 and 1930 anticipated and provided for the eventual collapse of this core of rock. In November 1936 the core settled down gradually on the edges of the stopes, descending a total of about 50 ft. vertically. The men in the stopes had ample warning and no lives were endangered. There were no air blasts in the mine. The core broke loose from its top support at the 500-ft. level along old fracture planes and evidently is crushed to the ninth level. Movement was of short duration; the following shift went to work in the stopes as usual. No ore was lost. Geological Structure The shape of the orebody is illustrated in Fig. 1. The geology of the Colorada orebody has been described by V. D. Perry1 and by W. G.

Jan 1, 1940

By C. D. Dolman

SINCE the outbreak of the war we have discovered in the United States minerals of which there was no general knowledge, and which compared very favorably with anything that could be found in any foreign country. One of these was magnesite, or the carbonate of magnesium as it occurs in nature. This mineral crystallizes in the rhombohedral form, has a specific gravity of 3.00, molecular weight of 84 4, molecular volume of 28.1, and a hardness of 3.5 to 4.5.1 Had it not been for the development of the California deposits of this mineral and the discovery of immense new deposits in Stevens County, Wash., the production of steel during the war would have been more seriously handicapped than it was. Now, that the war is over, we face the possibility of competition' with the cheap Austro-Hungarian magnesite; and while that may not he so cheap as it was before the war, adequate protection should be given the producers in this country to insure the full development of our own resources. The consumption of magnesite in the United States in 1913, the year preceding the war, was 172,591 tons of calcined and 22,872 tons of crude magnesite. Less than 3 per cent. of this amount was produced at home. At the present time, practically all of the magnesite used in the United States is produced at home. Expensive plants have been constructed, extensive exploration work done, and production pushed to the utmost to supply the necessary requirements of the war. Thus a considerable industry has been built up within our borders. Some magnesite has been imported from Canada and Mexico, but the Canadian ore was of an inferior grade and could not be used satisfactorily. During the year 1918, production decreased in both the United States and Canada; the production in the United States was 225,000 tons in 1918 as against 315,000 in 1917, and in Canada 39,365 tons in 1918 as against 58,090 in 1917.

Jan 8, 1919

By A. F. Crawley

THIS communication reports measurements of the densities of dilute liquid alloys of 0.1 and 0.5 pct Ag in Sn over temperature ranges of 170" and 210°C, respectively, above the melting point of pure tin (232°C). The work was earried out to provide accurate density data for some liquid diffusion studies performed by Davis.1 Densities were measured using a pycnometric technique established in this laboratory. Earlier papers2, 4 attest to the accuracy of ±0.05 pct and the reliability of the method. Tin, 99.999 pct pure, and silver 99.99 pct pure, were melted in a graphite crucible in a vacuum furnace and held at temperature for 10 hr to ensure homogenization. Ingots suitable for machining

Jan 1, 1970

By P. L. De Bruyn

The adsorption density of dodecylammonium ions at the quartz-solution interface has been Theadsorptiondensitydetermined as a function of collector concentration and pH. A ten thoushasbeenandfold range of amine salt concentration was covered at neutral pH. Experimental results show that over a thousandfold concentration range at neutral pH, the adsorption density (I) is proportional to the square root of collector concentration. Except at high concentrations, I increases with increases with increasing pH, but in general this effect is surprisingly small. . , . . A critical pH curve has been established for the flotation of quartz with dodecylammonium acetate. The conditions along the flotation curve are correlated with the adsorption measurements. THE behavior of collectors at the mineral-solution interfaces is usually explained in terms of an ionic adsorption process. Through the distribution of collector ions between the solid surface and the- co-existing solution phase the mineral is believed to acquire a water-repellent surface coating. Quantitative adsorption studies have been made on simple flotation systems1-4 only within the last few years. Such investigations were made possible by the adoption of the radiotracer method of analysis. As a consequence of these studies a new parameter has been added to aid the understanding of the flotation process. The research investigation to be discussed in this paper was undertaken to obtain a better understanding of the behavior of a cationic-type collector. This objective was approached through the determination of the distribution of dodecylammonium acetate between the quartz-solution interface and the solution as a function of the collector salt concentration and pH. To bring this investigation to focus on the more practical aspect of flotation research, an attempt was also made to correlate the adsorption results with actual flotation tests. Quartz: A —100 mesh ground crystalline quartz was infrasized; the products of the third and fourth cones were mixed together and reserved for experimental purposes. This stock material was cleaned by leaching in boiling concentrated HC1. After leaching the quartz was rinsed with distilled water until the filtrate showed no trace of chloride ian. It was then washed several times and dried. The qwrtz had a specific surface of 1400 cma per g as deterhined by the krypton gas adsorption method. Collector: The distribution of dodecylammonium acetate between the quartz surface and the solution phase was determined by the radiotracer method of analysis with carbon 14 as the tracer element. The radioactive amine salt with C" synthesized into the hydrocarbon chain5 was supplied by Armour and Co. The tracer element was located adjacent to the polar group. The radioactive salt as received had a specific activity of about 0.14 mc per g. When desired, dilution of this activity was effected by addition of non-radioactive dodecylammonium acetate also supplied by Armour and Co. ........ All other inorganic reagents used in this research were of reagent grade. Conductivity water was used for making up all solutions. Adsorption Tests: Two different experimental methods were used. In the first, to be designated as the agitation method, a weighed amount of quartz and a measured volume of amine salt solution were agitated in a 100-ml or 50-ml glass-stoppered pyrex graduated cylinder. The cylinder was filled with solution up to the stopper, since erratic results were obtained when an air space was left over the suspension. Time of agitation varied from 1 to 2 hr. Preliminary tests at different agitation times showed that the amount adsorbed remained constant for all agitation times exceeding 1/2 hr. After this conditioning period, the solids were separated from the solution by filtration through a Buechner fritted-disk funnel. The solution was re-circulated 10 times or more to allow the fused silica disk to come to equilibrium with it. Determinations of the amount of amine adsorbed on the frit itself indicated that this amount was less than 10 pct by weight of the amine acetate abstracted by 10 g of quartz. The funnel with quartz covered by a thin layer of solution was then centrifuged for approximately 5 min, at which time the moisture content of the solids was reduced to about 5 pct by weight. The wet quartz was blown into a tared beaker, re-weighed and allowed to dry at room temperature. A final weighing was then made to determine the moisture content. The second experimental method, similar to the procedure adopted by Gaudinand Bloecher,' will be referred to as the column method. Two liters of solution were passed through a bed of quartz contained in a Buechner funnel attached to a pyrex separatory funnel by means of a ball and socket joint. Preliminary tests showed that increasing the volume of solution above 2 liters does not give a measurable increase in adsorption. From 4 to 4 1/2 hr were required for 2 liters of solution to pass through the column. The moisture content of the quartz was again reduced to a minimum by centrifuging. A slightly modified column apparatus was used for experimenting with alkaline amine solutions. The same basic unit was used, but the underflow from the Buechner funnel was again fed into a Separafory

Jan 1, 1956

By E. P. Shoub, J. P. Riott, R. C. Buehl

Pilot-plant tests have demonstrated that it is possible to produce low-sulphur sponge iron (0.03 to 0.0; per cent sulphur) as a continuous process in an internally fired rotary kiln from iron ore or mill scale; a solid carbonaceous reducing agent; and dolomite to control the sulphur content. The proportion of dolomite employed is small, so that it does not add materially to the cost of the sponge iron. The entire operation is carried out at temperatures well below the melting point of iron, so that the product is discharged in granular form. Compared with other methods for the production of sponge iron, the rotary-kiln method has the advantage of simplicity and low cost of equipment and, in addition, requires little labor for the operation of the process. The pilot-plant equipment used for this work was sufficiently large to demonstrate that the production of low-sulphur sponge iron in rotary kilns is feasible and that the costs are not excessive for applications in which sponge iron has distinct advantages over competing materials. This report constitutes a part of a program by the Bureau of Mines under a special Congressional appropriation for studying gaseous and solid fuel reduction of iron ores. Numerous articles on various phases of these investigations have appeared in Bureau of Mines publications or technical journals. Other articles are being prepared and will be published in the near future. The present program, which is under the supervision of Dr. R. S. Dean, Assistant Director of the Bureau of Mines, was initiated in 1942. However, a considerable portion of the work was a continuation of previous research performed by the Bureau. Previous investigators1-4 utilizing the rotary kiln for the production of sponge iron did not develop a suitable method for controlling the sulphur content. A review of the literature shows that the sulphur content in the product usually exceeded 0.1 per cent unless reducing agents with very low sulphur contents were employed. This high sulphur value made the product generally undesirable for use as melting stock in electric or open-hearth furnaces. The production of sponge iron in an internally fired rotary kiln has been described in detail in Bureau of Mines Bullelin 2702; consequently, this article will be confined mostly to new features of the process—that is, means for controlling the sulphur content of the sponge iron. Suitable raw materials include iron ores or mill scale; coal or coke for reducing them, and sized dolomite ( — poi-100-mesh) as the sulphur-controlling agent. A mixture of the raw materials in granular form is fed into the elevated end of the kiln, which is fired axially from the lower or discharge end. In the reducing zone, which is approximately one third the length of the kiln measured from the dis-

Jan 1, 1948

By R. C. Buehl, J. P. Riott, E. P. Shoub

Pilot-plant tests have demonstrated that it is possible to produce low-sulphur sponge iron (0.03 to 0.0; per cent sulphur) as a continuous process in an internally fired rotary kiln from iron ore or mill scale; a solid carbonaceous reducing agent; and dolomite to control the sulphur content. The proportion of dolomite employed is small, so that it does not add materially to the cost of the sponge iron. The entire operation is carried out at temperatures well below the melting point of iron, so that the product is discharged in granular form. Compared with other methods for the production of sponge iron, the rotary-kiln method has the advantage of simplicity and low cost of equipment and, in addition, requires little labor for the operation of the process. The pilot-plant equipment used for this work was sufficiently large to demonstrate that the production of low-sulphur sponge iron in rotary kilns is feasible and that the costs are not excessive for applications in which sponge iron has distinct advantages over competing materials. This report constitutes a part of a program by the Bureau of Mines under a special Congressional appropriation for studying gaseous and solid fuel reduction of iron ores. Numerous articles on various phases of these investigations have appeared in Bureau of Mines publications or technical journals. Other articles are being prepared and will be published in the near future. The present program, which is under the supervision of Dr. R. S. Dean, Assistant Director of the Bureau of Mines, was initiated in 1942. However, a considerable portion of the work was a continuation of previous research performed by the Bureau. Previous investigators1-4 utilizing the rotary kiln for the production of sponge iron did not develop a suitable method for controlling the sulphur content. A review of the literature shows that the sulphur content in the product usually exceeded 0.1 per cent unless reducing agents with very low sulphur contents were employed. This high sulphur value made the product generally undesirable for use as melting stock in electric or open-hearth furnaces. The production of sponge iron in an internally fired rotary kiln has been described in detail in Bureau of Mines Bullelin 2702; consequently, this article will be confined mostly to new features of the process—that is, means for controlling the sulphur content of the sponge iron. Suitable raw materials include iron ores or mill scale; coal or coke for reducing them, and sized dolomite ( — poi-100-mesh) as the sulphur-controlling agent. A mixture of the raw materials in granular form is fed into the elevated end of the kiln, which is fired axially from the lower or discharge end. In the reducing zone, which is approximately one third the length of the kiln measured from the dis-

Jan 1, 1948

By D. H. Davis

HAND picking was the earliest form of coal preparation, first practiced to improve the outward appearance of the coal being loaded and to remove any pieces that might appear objectionable to the buyer. It is still extensively practiced on the plus 1-in. sizes and particularly on the plus 4-in. and larger sizes. The percentage of ash reduction that can be obtained by hand picking depends entirely upon the amount of free impurity occurring in the coal. In a particularly dirty seam, where a considerable quantity of rock occurs with the coal as mined, hand picking produces sizable reductions in ash. If the coal is prepared clean at the working face hand picking is a final check; in some cases, on the smaller sizes, it is practiced only to show that an effort is being made to prepare a clean product. Because of the variety and different methods of handling coal, the early plants were not always adapted for e5cient hand picking. At most mines an attempt is made to - remove the visual impurities at one or more of the following places in the surface plant: (1) in the tipple during screening; (2) on picking band, belt, or shaking conveyors; (3) during the loading of the coal as it goes into the railroad car. Whatever the physical arrangements of the plant may be, the usual practice is to run the coal over the screen, picking table or boom at such a rate that each piece of impurity is exposed to the view of the picker. Coal should be uniformly sized and spread out on the table so as not to exceed one particle in depth. The flow should be even in tonnage and the speed of travel should be from 30 to 60 ft. per minute and not exceed 80 ft. per minute. The dirtier the coal, the slower the speed required for a fixed number of pickers. Actually, picking capacity depends on the number of particles per minute that can be picked out and rejected. The percentage of impurity removed is determined by this and the average weight of piece rejected. Handling a mixture of sizes from nut to the largest lump over the same picking table or conveyor is not conducive to efficient separation by hand picking because the lump hides the smaller impurities. Picking tables should not exceed 5 ft. in width, so that a picker on either side will be able to remove impurities from the center of the table. Guards usually are placed at the side of the screen, boom or table, to allow the pickers to stand close enough to remove the impurities without too long a reach.

Jan 1, 1950

By R. S. Busk, E. G. Bobalek

THE relation between gases and metals has been a subject of increasingly active investigation during the past years, principally devoted to the study of metal-hydrogen systems. It has been found that hydrogen is soluble in most molten metals to an appreciable extent and somewhat soluble in most solid metals.1,2 In every known case there is a sudden decrease in solubility at the melting point. Thus a freezing mass of metal containing hydrogen in excess of the solid solubility must in some way expel that excess during the freezing process. This leads to internal blowholes, and such effects as pinhole porosity. Gas trapped in metal that is to be fabricated by a working process such as extrusion or rolling may be the cause of blisters. For some cases, notably in copper and iron alloys, the presence of hydrogen can lead to extreme embrittlement of the metal. The aluminum-casting industry has long recognized the importance of hydrogen in its techniques. Excess gas in molten aluminum alloys often leads to the formation of "pinhole porosity" in the solid casting.3 To prevent this hydrogen pickup it is necessary to avoid overheating the melt and to protect it from such sources of hydrogen as water. In spite of all precautions, it is sometimes necessary to extract hydrogen from the melt just before pouring by bubbling with a gas such as chlorine or nitrogen. Recently it has been found that somewhat similar effects are possible with magnesium alloys, in that the presence of hydrogen can induce or aggravate microporosity.4-6 This paper is primarily concerned with a presentation of some facts concerning the solubility of hydrogen in magnesium and its alloys, with some discussion of the significance of these facts in a practical sense. In any study of gas-metal systems it is important to bear in mind a clear distinction between adsorption on surfaces and absorption within the lattice.1 Considerable attention has been directed toward determining whether hydrogen is truly soluble in any solid metal, or whether the gas extracted from a sample was present at surfaces (external and internal) only.7 Most present day evidence supports the idea of true solution in addition to adsorption. A second important factor when dealing with gas solubility is the dependence of such solubility on external pressure. For diatomic gases such as H2, solubility in both liquid and solid metals apparently is proportional to the square root of the pressure, indicating a breakdown to a monatomic gas before diffusion into the lattice .2 For true equilibrium values of solubility, therefore, the pressure at the metal surface must be constant and, preferably, known. Very little has been published concerning the hydrogen-magnesium system. Winterhager has presented the most

Jan 1, 1946

By G. R. Countryman, G. H. Thomas, I. Fatt

Miscible displacement in both thewetting and non-wetting phase has been studied in two-phase systems. Experimental data show that dispersion is a function of saturation and cannot be predicted from dispersion theory developed for single-phase miscible displacement. Some of the effect of saturation on dispersion is a result of the development of dead-end pores or a dendritic pore structure in multiphase systems. The presence of dead-end pores or dendritic structure is confirmed by pressure transient studies. INTRODUCTION Displacement by flooding with a miscible liquid is a possible means for recovering the estimated two-thirds of the oil that remains behind after primary production. The immense economic importance of a process that can recover such large quantities of oil has led to extensive laboratory studies of miscible displacement. As usual in production research, most laboratory studies of miscible displacement have not attemped to reproduce all of the conditions existing in a petroleum reservoir. In the early stages of investigation the need has been for information on the broad, basic principles of the phenomena involved in miscible displacement. The effects of overburden pressure, reservoir temperature, and wettability have been considered of secondary importance. However, one property of a petroleum reservoir which is expected to be of major importance and yet has been omitted from many laboratory studies is the presence of interstitial water. Two possible effects of interstitial water on the displacement mechanism in the hydrocarbon phase immediately come to mind. First, from the general1y accepted theory that capillarity governs the distribution of oil and water in porous rock one would expect that for water-wet rock the water will be in the small pores and oil in the large pores. A miscible displacement of oil carried out in the presence of water is operating in a pore size distribution shown in Fig. lb, whereas if the same test had been performed with only one phase present the pore size distribution is as shown in Fig. la. Although there is not yet a theory of miscible displacement which explains in detail the effect of pore size distribution, one would expect the differences between Figs. la and lb to influence the displacement efficiency. A second factor which may make a multiphase system different from a single-phase system is the presence in the multiphase system of dead-end pores or dendritic structure. Experiments of various kinds on reservoir rock have led to the belief that all pores in the network structure of a porous rock take part in conducting fluid during single phase fluid flow.1,2 There are then no dead-end pores and no fingers or dendritic structures containing stagnant fluid. In a multiphase system, however, the second phase may

By J. G. Paterson, J. A. Meech

Abstract-Activated carbon is a strong adsorbent for amyl xanthate, capable of removing from solution up to a quarter of its own weight in xanthate. In selective flotation systems where depression is unstable, separation efficiency increases with activated carbon additions. The improvements observed are brought about by two mechanisms. First, by removing stripped and residual collector ions from solution, competition between depressant and xanthate is reduced and depression is stabilized. In addition, the texture of the froth phase at high pine oil concentrations is modified by activated carbon such that better drainage between the air bubbles takes place, alleviating the mechanical entrapment of depressed mineral particles. In treating a bulk copper/lead concentrate produced from GECO ore, selectively between galena and chalcopyrite is increased when activated carbon is used. Although the rate of flotation of galena is marginally lowered by activated carbon, a small extension in the duration of flotation allows the lead recovery to increase by 19% at the same recovery of copper. An addition rate of 0.75 kg/ton (1.5 lb/st) of bulk concentrate is suggested as optimum from this testwork. However, the actual level wed in the mill is dependent upon the residual concentrations of collector and frother in the bulk concentrate Pulp.

Jan 11, 1978

By Francis B. Foley

Fundamentally, slag control in acid open-hearth steelmaking cannot be considered from any different viewpoint than can that in basic open-hearth melting. The same laws of mass action and distribution apply in both cases, so that while most, if not all, of the recent intensive study of slag-metal reactions have been carried on in basic furnaces the principles evolved are nevertheless applicable to acid open-hearth steelmaking and it is therefore not our intention to discuss the various possible chemical reactions in the slag and between slag and metal. The points we have to carry in mind, in considering acid melting practice, are the differences in the type of steels produced and the difference in the chemistry of both the steel and the slag. The acid furnace is used to produce, let us not say "quality" steel, for, after all, a "quality" steel for one purpose is not a "quality" steel for another, but steels that require acid melting practice to be of proper "quality." Steels that are required to have high physical properties in tests taken transversely from forgings are peculiarly the product of acid furnaces. In this class may be placed various ordnance materials, notably gun forgings and projectiles and large commercial forgings requiring high transverse ductility. Forgings requiring an extremely hard surface, polished and free from imperfections, are also typical acid-steel products. Of course, such forgings can be made of basic open-hearth steel, but not with the same assurance of success. At the start, the preparation of the slag in the acid and basic furnaces differs. Lime is charged in the beginning of a basic heat to form a slag, but in the acid furnace nothing is added for the special purpose of producing a slag when the heat is melted down. The result is that the acid furnace has much less volume of slag than the basic. The slag formed upon melting down is made up almost entirely of the oxides of elements contained in the metallic charge, notably oxides of iron, manganese and silicon, the amount of silica being augmented somewhat by erosion of the furnace lining. A typical acid open-hearth slag, when the bath is completely melted, will have an analysis about as follows: CaO, 3.40 per cent; SiO2, 51.9; Fe2O3, 1.25 ; FeO, 28.86; Al2o3, 1.82; MnO, 12.22. The high iron oxide content of the slag gives it a black appearance. The metal is high in carbon and some manganese and silicon will be present, depending on the amount in the original charge, the amount of

Jan 1, 1935

By Francis B. Foley

Fundamentally, slag control in acid open-hearth steelmaking cannot be considered from any different viewpoint than can that in basic open-hearth melting. The same laws of mass action and distribution apply in both cases, so that while most, if not all, of the recent intensive study of slag-metal reactions have been carried on in basic furnaces the principles evolved are nevertheless applicable to acid open-hearth steelmaking and it is therefore not our intention to discuss the various possible chemical reactions in the slag and between slag and metal. The points we have to carry in mind, in considering acid melting practice, are the differences in the type of steels produced and the difference in the chemistry of both the steel and the slag. The acid furnace is used to produce, let us not say "quality" steel, for, after all, a "quality" steel for one purpose is not a "quality" steel for another, but steels that require acid melting practice to be of proper "quality." Steels that are required to have high physical properties in tests taken transversely from forgings are peculiarly the product of acid furnaces. In this class may be placed various ordnance materials, notably gun forgings and projectiles and large commercial forgings requiring high transverse ductility. Forgings requiring an extremely hard surface, polished and free from imperfections, are also typical acid-steel products. Of course, such forgings can be made of basic open-hearth steel, but not with the same assurance of success. At the start, the preparation of the slag in the acid and basic furnaces differs. Lime is charged in the beginning of a basic heat to form a slag, but in the acid furnace nothing is added for the special purpose of producing a slag when the heat is melted down. The result is that the acid furnace has much less volume of slag than the basic. The slag formed upon melting down is made up almost entirely of the oxides of elements contained in the metallic charge, notably oxides of iron, manganese and silicon, the amount of silica being augmented somewhat by erosion of the furnace lining. A typical acid open-hearth slag, when the bath is completely melted, will have an analysis about as follows: CaO, 3.40 per cent; SiO2, 51.9; Fe2O3, 1.25 ; FeO, 28.86; Al2o3, 1.82; MnO, 12.22. The high iron oxide content of the slag gives it a black appearance. The metal is high in carbon and some manganese and silicon will be present, depending on the amount in the original charge, the amount of

Jan 1, 1935

By S. Feuerstein, J. M. Galligan

SURFACE effects in the deformation of metal single crystals have been noted by a variety of workers.&apos; A large majority of these experiments have used surface roughening or a second chemical constituent at the surface. We performed the following experiment to see how the presence of a fine polycrystalline layer influenced the stress-strain behavior of single crystals of zinc. Spherical single crystals oriented for single slip were acid-machined into cylindrical dumbbell-shape shear specimens having a gage length of 3.2 mm and a cross section of 20 sq mm. The crystals were then deformed varying amounts at the surface of the gage section with emery paper. Following this treatment, the crystals were heated for approximately 8 hr at 200°C. This resulted in a polycrystalline surface layer, with an observed grain size of approximately 1/10 mm in diameter. Deformation of these crystals was undertaken at 78°K in pure shear, the shear direction <1120> lying in the basal plane. The deformation apparatus has been described elsewhere.2 The presence of a polycrystalline layer markedly raises the first observed plastic deformation, as can be seen from Fig. 1. This yield stress is then followed by a region of low work hardening comparable to that observed in shear tests on similar untreated zinc single crystals. However, the extent of easy glide is quite seriously curtailed, and is followed by a more rapid increase in work hardening, reminiscent of Stage II observed in the deformation of fcc crystals. The higher yield can probably be attributed to the difficulty of penetrating the grain boundaries by dis locations from sources interior to the boundaries. If we associate this yield stress with this breakthrough, then we might expect that more dislocations would be trapped in the crystal than is normally the case. If there is an increased density of trapped dislocations, this could lead to an increase in the amount of trapping with increasing strain, giving a larger strain hardening in the crystal. Although other explanations can be given, it is quite tempting to characterize the region of increased work hardening as a manifestation of Stage II-type deformation. This seems reasonable if the polycrystalline layer is quite efficient in holding up a large number of dislocations. In zinc crystals, since it is well-known that the slip distance can be as large as the specimen dimensions, one would only expect to entrap a large number of dislocation with very large crystals or by applying surface layers. If this interpretation of increased strain hardening is correct, it might present a method of studying Stage II in single crystals of zinc. This research was supported, in part, by the Office of Naval Research (1959). The encouragement and stimulation given by Professor E. R. Parker is gratefully acknowledged.

Jan 1, 1965

By Earl A. Gulbransen, Kenneth F. Andrew

DRY oxidation of zirconium has been studied by several groups.&apos;" The present work extends our early study1 to the high-temperature studies of Cubicciotti2 and Belle and Mallett.8 ulbransen and Andrew1 showed that the parabolic rate law fitted the 2-hr experiments between 200" and 425°C after an initial deviation. An energy of activation of 18,200 cal per mol was calculated. Cubicciotti,&apos; in a study between 593" and 880°C, showed that the parabolic rate law fitted the data and an energy of activation of 32,000 cal per mol was calculated. In contrast to these studies on sheet specimens, Belle and Mallett3 studied the oxidation reaction on rod specimens and found a cubic rate law to fit the data.

Jan 1, 1958