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By W. W. Yankie, A. T. Chatas

This paper is written as a discussion of the paper, "Buckling of Tubing in Pumping Wells, Its Effects and Means for Controlling It" by Arthur Lubinski and K. A. Blenkarn, which war published in the March, 1957, issue of Journal oP Petroleum Technology. An analogy is drawn between the buckling of tubing in pumping wells and a fiube, fitted with frictionless pistom, linked by an inelastic rod, and which is subject to an internal pressure. A simple derivation shows that the critical internal pressure exerts a hydrostatic force on the pisfon which is close to the Euler load for compression buckling of the same tube. While the results agree with those of Ref. I, the basis of calculations has been changed and clarified. The presence of internal upsets at the pistons is discussed. INTRODUCTION In the Appendix, Lubinski and Blenkarn state that a tube with ends closed by frietionless pistons connected by an inelastic rod (Fig. 1) will behave as a column loaded with a force equal to the tension force in the rod. They note that, while the fluid may not resist a shear force, the moment in every cross section is piston force multiplied by the elastic line deflection (c.f., the loaded column). Since the equations of the elastic lines are similar in the two cases, the solution is carried over from the loaded column. On the contrary it is suggested that while the method of solution remains unchanged from the column case, the actual stress distribution is different, a fact which is taken into account in this note. A simple method of approximating the buckling pressure is given, and their reasoning in our mathematical language is continued. GENERAL EQUATIONS The equation of an elastic line, initially straight in the x direction (Fig. 2), is given by In the case of a loaded column, this may be written: For the pressurized tube it was shown, by analogy with the loaded column, that tion has a solution that, for P, = EI/P, any deflection is stable (a criterion for buckling). An attempt will be made to describe more precisely, the forces acting on the tube, and to analyze the buckling condi,tions. DISCUSSION If the tube assumed a deflected position, the volume between the pistons would increase so that in deflecting, the energy of the fluid would decrease. When the increase in elastic energy of the tube (due to bending) is equal to the decrease of fluid pressure energy, buckling conditions exist. Analyzing the forces on the tube (Fig. 3), two transverse forces, T,, T,, exerted by the pistons, are balanced by a resultant of the pressure elements which are linearly dependent on the local curvature of the elastic line (Fig. 4). Unlike most buckling cases, the forces depend on the shape of the elastic line, a condition that makes a rigorous mathematical treatment difficult. Several approximations may be used to clarify the problem.

By J. Intrater, E. S. Machlin

In a study of grain boundary sliding in superpure Copper (99.999 pct),&apos; bicrystals of a specific shape were used, Fig. 1. That shape was adopted in order to minimize the deformation of grains on both sides of the boundary due to clamping. Another requirement was to obtain bicrystals with a rnicroscopically straight boundary, Fig. 2. Single crystals of predetermined orientation (growth axis parallel to <310>) were grown in a graphite mold from seed crystals. The apparatus for growing single crystals consisted of a vacuum system, argon purification train, vycor tube (in which a graphite mold was placed), a carriage for the induction coi1,and a Lepel (7-1/2 KW) high-frequency converter. The graphite mold was heated by induction and the charge in turn was heated by conduction. The induction coil was moved by means of a motor at a speed of 1 1/2 in. per hr. Single crystals of two different orientations were grown from the melt in one-half of a split graphite mold, Fig. 3(a). Two such crystals were then placed in contact with each other in another split mold, Fig. 3(b) using tapered inserts. The pair of crystals was then remelted starting 2 in. from the top to insure that the whole top portion of the crystals was maintained molten. By moving the induction coil

Jan 1, 1960

By D. Walton, W. C. Winegard, W. A. Tiller, J. W. Rutter

It is shown that the cellular substructure observed in binary alloy crystals grown from a melt of known concentration can be eliminated by the proper choice of growth conditions. For a given solute concentration, there exists a critical ratio of temperature gradient G in the melt to the rate of solidification R of the crystal, i.e., of G/R, which must be exceeded before the cellular structure can be eliminated. A secondary substructure is observed which is related to the cellular substructure. IN crystals grown from the melt, a prismatic substructure has been observed by Buerger,1 Pond and Kessler,2 and Rutter and Chalmers." This structure is columnar in the direction of growth and appears as a network of hexagonal cells in the plane of the solid-liquid interface. Rutter and Chalmers3 have shown that this substructure occurs as a result of the presence of a layer of high impurity concentration in the melt adjacent to the solid-liquid interface. The formation of the substructure is accompanied by a redistribution of this impurity in the neighborhood of the interface. The structure can be controlled by proper adjustment of the variables in the solidification process such as a rate of solidification and the temperature gradient in the liquid. The redistribution of solute during the solidification of metal crystals has been treated theoretically

Jan 1, 1956

By Fredrick C. Pearson, Albert Brown, Emil R. Ermert

EARLY in 1946 the Shen-Penn Production CO., a subsidiary of the Philadelphia and Reading Coal and Iron Co., was organized to operate the Shenan-doah Stripping, one of the largest open-pit anthracite mines in the world. This mine is located in the Western Middle Field near Shenandoah, Schuylkill County, Pa., and involves the removal of approximately 60,400,000 cu yd of overburden to permit the recovery of 9,200,000 gross tons of rough cleaned coal from an area 8000 ft long and 2700 ft wide, with a maximum depth of 400 ft. Operations began in August 1946, and within the next few years 31,811,-000 cu yd of overburden and 2,994,000 gross tons of rough cleaned coal had been removed. During the early days of open-pit mining operations, steam was the primary motive power. However, as the size of the equipment increased, so did the problems of fuel, water, labor, and maintenance. Because of its efficiency and convenience, electric power rapidly replaced steam. Later the modern diesel engine was developed, and today practically all stripping equipment is either electric, diesel-electric, or diesel-powered. In locations where electric power is not readily available or where operations are scattered over a large area, diesel power is favored. At Shenandoah operations are highly concentrated in a relatively small area, and this fact, coupled with the availability of ample electric energy, indicated the selection of electric-powered equipment. Since this was a new company, it was necessary to purchase every item, from the smallest tools to stripping shovels. Although the organization period was hampered by the postwar scarcity of materials, an excellent opportunity was afforded to set up a thoroughly modern plant. This paper will describe the power facilities, with particular emphasis on the more unusual features. The substation is a major item in a power distribution system, and a simple arrangement utilizing high quality equipment is an excellent investment. With this in mind, a 3000-kva unit type substation was selected and installed. This unit is made up of three sections, which are assembled simply by bolting together the various sections and connecting the power leads, see Fig. 1. The transformer proper, rated at 3000 kva, is oil-filled and nitrogen-sealed, having a 66,000-v Delta primary and 4160-v Wye secondary with a grounded neutral and 46-ohm ground resistor. The function of this resistor will be described later. Electrical energy is purchased from a public utility and is delivered to the main substation at 66,000 v over one or the other of two incoming lines. These lines terminate on a high voltage switching section on which high voltage fuses and Thyrite lightning arresters are installed. The outgoing feeder switching unit is made up of the following: 1—An auxiliary compartment consisting of potential and current transformers with necessary terminal blocks for connecting to power company metering equipment. 2—Another auxiliary compartment containing a control transformer, battery, and battery charger for the furnishing of power for tripping and reclosing of power circuit breakers. The assemblies in these auxiliary compartments are mounted on carriages that can be rolled in and out with ease. Potential transformers are readily accessible for inspection and fuse replacement. When withdrawn from the operating position they are completely de-energized and grounded. 3—Two feeder breaker compartments, each with a 1200-amp, 5000-v, solenoid-operated magneblast air circuit breaker, voltmeter, ammeter, overcurrent relays, and a ground relay. These power circuit breakers can be quickly and safely withdrawn or repaired. A built-in isolating mechanism is used for lowering the breaker to a transfer truck for repair or inspection or as a means of disconnecting the breaker from the source of power as an added precaution when working on power lines. Thus inspec-

Jan 1, 1953

By Albert Brown, Emil R. Ermert, Fredrick C. Pearson

EARLY in 1946 the Shen-Penn Production CO., a subsidiary of the Philadelphia and Reading Coal and Iron Co., was organized to operate the Shenan-doah Stripping, one of the largest open-pit anthracite mines in the world. This mine is located in the Western Middle Field near Shenandoah, Schuylkill County, Pa., and involves the removal of approximately 60,400,000 cu yd of overburden to permit the recovery of 9,200,000 gross tons of rough cleaned coal from an area 8000 ft long and 2700 ft wide, with a maximum depth of 400 ft. Operations began in August 1946, and within the next few years 31,811,-000 cu yd of overburden and 2,994,000 gross tons of rough cleaned coal had been removed. During the early days of open-pit mining operations, steam was the primary motive power. However, as the size of the equipment increased, so did the problems of fuel, water, labor, and maintenance. Because of its efficiency and convenience, electric power rapidly replaced steam. Later the modern diesel engine was developed, and today practically all stripping equipment is either electric, diesel-electric, or diesel-powered. In locations where electric power is not readily available or where operations are scattered over a large area, diesel power is favored. At Shenandoah operations are highly concentrated in a relatively small area, and this fact, coupled with the availability of ample electric energy, indicated the selection of electric-powered equipment. Since this was a new company, it was necessary to purchase every item, from the smallest tools to stripping shovels. Although the organization period was hampered by the postwar scarcity of materials, an excellent opportunity was afforded to set up a thoroughly modern plant. This paper will describe the power facilities, with particular emphasis on the more unusual features. The substation is a major item in a power distribution system, and a simple arrangement utilizing high quality equipment is an excellent investment. With this in mind, a 3000-kva unit type substation was selected and installed. This unit is made up of three sections, which are assembled simply by bolting together the various sections and connecting the power leads, see Fig. 1. The transformer proper, rated at 3000 kva, is oil-filled and nitrogen-sealed, having a 66,000-v Delta primary and 4160-v Wye secondary with a grounded neutral and 46-ohm ground resistor. The function of this resistor will be described later. Electrical energy is purchased from a public utility and is delivered to the main substation at 66,000 v over one or the other of two incoming lines. These lines terminate on a high voltage switching section on which high voltage fuses and Thyrite lightning arresters are installed. The outgoing feeder switching unit is made up of the following: 1—An auxiliary compartment consisting of potential and current transformers with necessary terminal blocks for connecting to power company metering equipment. 2—Another auxiliary compartment containing a control transformer, battery, and battery charger for the furnishing of power for tripping and reclosing of power circuit breakers. The assemblies in these auxiliary compartments are mounted on carriages that can be rolled in and out with ease. Potential transformers are readily accessible for inspection and fuse replacement. When withdrawn from the operating position they are completely de-energized and grounded. 3—Two feeder breaker compartments, each with a 1200-amp, 5000-v, solenoid-operated magneblast air circuit breaker, voltmeter, ammeter, overcurrent relays, and a ground relay. These power circuit breakers can be quickly and safely withdrawn or repaired. A built-in isolating mechanism is used for lowering the breaker to a transfer truck for repair or inspection or as a means of disconnecting the breaker from the source of power as an added precaution when working on power lines. Thus inspec-

Jan 1, 1953

By Kenneth N. Weaver

Portland cement can be made from relatively abundant industrial minerals and rocks, and this may explain why cement producers placed little emphasis on geology during the early days of the industry. After World War II, however, the industry began to recognize the need for geological exploration, and some of the larger companies began to build a permanent geological staff. This development had its roots in the following factors: Rapid expansion of the industry; closing of many cement plants due to inadequate raw materials; necessity for upgrading the quality of portland cement; and the increasingly large capital investment required to build new cement plants, which in turn, required a higher degree of accuracy in reserve calculations. Whereas the geologist&apos;s function during the active expansion of capacity was primarily in exploration, one of his major functions today is the development of existing reserves.

Jan 1, 1965

By S. V. Krishnan, I. Iwasaki

In the upgrading of finely disseminated iron ores, selective desliming is the critical step that must be controlled to achieve efficient flotation. A prerequisite for selective desliming is a properly-dispersed pulp; sodium silicate is commonly used as a dispersant. The mechanism by which sodium silicate acts as a dispersant in the presence of calcium ions was examined by streaming potential measurements, settling tests, abstraction density determinations, selective flocculation tests, and scanning electron microscope observations. The study established that calcium silicate precipitates on the mineral surfaces dispersed the pulp. Although sodium silicate was a nonselective dispersant, the addition of corn starch aided flocculation of goethite over quartz at low silicate additions, but became ineffective at higher silicate levels. A model is presented to explain the dispersant mechanism of sodium silicate in the presence of calcium ion for selective flocculation of iron ores.

Jan 1, 1983

By R. F. Krueger

Results are presented from tests of dynamic fluid-loss rates to cores from clay-gel water-base drilling fluids containing different commercial fluid-loss control agents (CMC, polyacrylate or smt,ch), organic viscosity reducers (quebracho and complex metal lignosulfonate) and oil at several different levels of concentration. In the dynamic system the most effective individual additives to the clay-gel drilling fluid, based on cost-equalized concentrutiom, were found to be starch and the viscosity reducers. These results do not conform with the rankings determined by API fluid-loss rests, which indicate CMC, polyacrylate and starch to be the most effective and comparable. Generally, minimum dynamic fluid-losr rates were attained at cost-equalized concentrations of additive (including thinner) of about $1.00/ bbl, or less. For chernically treated clay-gel drilling fluids, both the standard and the high-pressure API filter-loss tests were found to he inaccurate indicators of trends in dynamic fluid-loss rates under the test conditions used, particulurly for drilling muds containing viscosity reducers. From a practical field viewpoint, restrictions on the applicability of the API fluid-loss test are such that it is open to question whether or not results of this test can be used routinely with confidence as an indicator of control of down-hole fluid loss under field treating conditions. INTRODUCTION The petroleum industry spends large sums of money during drilling operations to control the fluid-loss properties of drilling fluids based on the standard API filter-loss test,&apos; which is a static filtration system. Laboratory studies&apos; &apos; of dynamic filtration have shown that in a given time period filtrate loss from a circulating mud stream is greater than from a static system and that it is a function of linear mud velocity, pressure and the properties of the drilling fluid. Ferguson and Klotz&apos; and Horner, et al," observed that (I) the dynamic fluid-loss rates for the drilling fluids used were not related to the extrapolated API filter loss and (2) the drilling fluids with the lowest API filter losses did not have the lowest dynamic fluid-loss rates. However, there has been no published information on the relative effects on dynamic fluid-loss rate as a given drilling fluid is treated with increasing amounts of chemical additive to reduce the API filter loss. Such information is economically important because drilling-fluid costs rise rapidly as chemical requirements increase. This paper presents the results of a study of dynamic filtratioi rates to cores from a clay-gel water-base drilling fluid treated with various commercial viscosity reducers and chemical fluid-loss control agents. The dynamic fluid-. loss rates to cores are compared with the standard API filter-loss values at several different levels of additive concentration. Dynamic filtration rates were obtained in each experiment under two different simulated wellbore conditions: (1) filtration just above the bit through a new mud cake laid down dynamically on a freshly drilled formation and (2) filtration up-hole through a mud cake formed by deposition of a static filter cake on top of the initial dynamically formed cake. The latter case corresponds to the bottom-hole conditions existing above the bit when mud circulation is restarted after a stand of pipe has been added or a round trip has been made to change the bit. Except for the short-duration, high-rate filtration beneath the bit where no mud cake can form, these two conditions probably represent the two extremes of dynamic filtration. Because thickness of a dynamic mud cake formed on freshly exposed formation is limited by the shearing action of the mud stream, the filtration rate for this condition is high. On the other hand, once circulation is stopped and a static mud cake forms on top of the dynamic cake, re-starting circulation has only a small effect on the cake properties and filtration rate is much lower thereafter. A discussion of the mechanics of mud-cake deposition and dynamic filtration is outside the scope of this paper but may be found in more detail in publications by prior investigators. APPARATUS AND EXPERIMENTAL CONDITIONS The test equipment used to simulate the dynamic flow conditions existing during drilling was a modification of that described previously by Krueger and Vogel: A schematic flow diagram is shown in Fig. 1. In general, a power-driven, high-pressure mud pump capable of delivering up to 60 gallmin was used to circulate drilling fluid parallel to the faces of 1-in. diameter sandstone cores mounted in a 2 3/4-in. ID high-pressure test cell. Pump rates were controlled by means of a magnetic clutch to maintain an average axial fluid velocity of 110 ft/min in the annular space between the cell wall and a 1 1/2-in. rod positioned on the center line of the cell. The core specimens were Berea sandstone plugs sealed with plastic inside 1 1/8-in. OD tubes and were fluid-saturated prior to use. Burettes were used to accumulate fluid discharged from the cores. The mud sump shown was used for treatment and storage of the drilling-fluid samples during a particular test. The valve arrangement permitted either (1) circulating drilling fluid through the by-pass line while treating with

By A. W. Schlechten, W. J. Kroll

As this paper is written, the only method for the commercial production of lithium metal is by the fusion electrolysis of LiCl-KCl mixtures, as first proposed by Guntz.2 The details of the industrial process have not been made public but Osborg8 stated that an efficiency of more than go pct is obtained with a lithium metal recovery above 95 pct and that the metal is 99.5 pct pure. Pletenev9 gave a power consumption of 66 kw-hr per kilogram of lithium metal produced with a salt consumption of 9 kg of LiCl and 0.4 kg of KCI. Pletenev&apos;s figures indicate a somewhat lower recovery than that given by Osborg. In a recent report on German lithium plants, Motock7 gives figures that are also less favorable than those reported by Osborg. At the German plants the lithium recovery in electrolysis was only 83.4 pct, with a power consumption of too kw-hr per kilogram of lithium. The metal averaged only 97 pct pure, with sodium and potassium as the chief impurities. The presence of SO2, SO3, Si02, Ba, Ca, and appreciable amounts of Na and Fe2O3 were especially disturbing in the electrolysis, and it was necessary to run the cell for some time to eliminate most of these impurities and condition the bath. It may be noted here that most of these impurities do not interfere with a thermal reduction process. These facts show that the present method of producing lithium is fairly efficient and would suggest that other methods could not compete. However, fusion electrolysis has certain drawbacks, such as the necessity of supplying costly, low-voltage direct current and of collecting and processing the anodic chlorine if it is not to be wasted. An even greater disadvantage is that lithium chloride must be used for an electrolyte. It is one of the most expensive salts of lithium, since it cannot be produced in anhydrous form by precipitation from aqueous solution, as is done with LiF or Li2CO3. The reduction of lithium by a vacuum process similar to the ferrosilicon method for the production of magnesium would have certain advantages. Small units, such as Pidgeon retorts, would give flexibility of production to meet demand, and a small plant would require moderate capital outlay. There is a possibility of using idle Government-owned magnesium plants for such purposes. The raw materials needed for vacuum thermal reduction would not have to be extremely pure, especially with respect to the iron and calcium content. This paper describes a series of experiments to determine the effectiveness of various agents for the reduction of lithium compounds. The investigation is part of a program of vacuum metallurgy con-

Jan 1, 1947

By Richard Addison, Ronald O. Cranor

Tailings dam construction and operation of Rosario Dominicana&apos;s gold/silver mine in the Dominican Republic is described. Impoundment of 8.5 kt/d of cyanide-process tailings is required in an intensive agricultural area which is also an earthquake and high rainfall zone. The earthfill embankment dam, incorporating internal gravel drains, is described together with methods of construction used. Methods of tailings transport, solution reclaim, and runoff water diversion that have evolved are discussed.

Jan 1, 1982

Jan 1, 1957

(The following list contains the names of those members of the Institute of whose connection with military service we have only recently become acquainted; it also includes the navies of a few who have recently been promoted or transferred. A complete list was published in the Year Book, issued as a supplement of the Bulletin for March, 1918.) ARNOLD, RALPH, Member of. Board of Tax Reviewers in connection with the administration of the War Revenue Act. BACON, RAYMOND F., Lieutenant-Colonel, Chemical Service Section, N. A., A. E. F. BALLENBERG, A. G., 1st Lieutenant, Ordnance Dept., U. S. A. BARCUS, WALTER J. E., Captain, Gun Division, Ordnance Officers&apos; Reserve Corps. BECKWITH, H. T., Captain, Engineer Reserve Corps, Headquarters Co., 312th Engineers.

Jan 6, 1918

By N. Marusov, D. K. Lowe, J. C. Karp

This paper considers, from an engineering viewpoint, several factors involved in creating, designing and locating horizontal barriers for controlling water coning. This is an effort to consolidate new concepts with previous information so that a reasonable selection of barrier materials, dimensions and vertical position can be made. Coning theories previously developed are briefly reviewed and an effort i,s made to reduce the results of coning-theory calculations to a point where routine calculations can be made with a desk calculator. It is expected that these simplified calculations will give usable predictions of the amounts of improvement attainable with barriers of various dimensions. Apparatus and procedures used for testing the suitability of commercial cements are described and test results are presented. INTRODUCTION Just about as soon as the phenomenon of water coning was recognized as an oilfield problem, there were suggestions that the production of water coning could be controlled or completely suppressed by means of horizontal barriers. It was also suggested that natural barriers such as shale streaks were helpful in restricting bottom-water production. The implication was that wells not having the benefit of continuous shale streaks to suppress bottom waters should, in some way, be supplied with an artificial barrier. With regard to the selection of barrier materials, there are many references which give the merits of liquid barriers.&apos;,&apos; These liquids include surface-active agents," recipitates&apos; and emu1sions. There is no real need to review the merits of these materials again. This paper deals with the use of solid barriers which could be made from various cements. The two major restrictions placed on these barrier cements are that (1) they must be commercially available in large quantities, and (2) they must be applicable by means of standard tools and equipment with techniques that have been fully developed and are in general use at the present time. A schematic drawing of a horizontal barrier in a reservoir is shown in Fig. 1. Two maximum stable water cones are illustrated-—-one for the wellbore diameter and the other for a barrier. A maximum stable cone is one that has attained its maximum volume at the critical production rate and is on the threshold of producing water. The volume displaced by each water cone and the rise in water table are indications of the amount of water-free oil produced. Corrections for porosity, connate water and residual oil, of course, must be applied. The water-oil interface which defines a maximum stable cone is called a free surface. It is a surface over which the pressure is constant. This free surface has the shape of a limiting streamline at and below which there is no movement or transport. The shape of the water cone describes the pressure drop in the reservoir. The effect of a completely impermeable barrier on the cone shape is essentially the same as extending the wellbore out to the barrier radius. The placement of a horizontal barrier requires that a horizontal fracture be created using a single-point entry technique. Two methods may be employed for filling this fracture. One method consists of propping the fracture before filling with cement. The other consists of fracturing with fracture fluid, then immediately following with cement

By David A. Summers, Clark R. Barker, Marian Mazurkiewicz

In May 1975 the U.S. Bureau of Mines contracted with the University of Missouri-Rolla, Rock Mechanics and Explosives Research Center to develop a Hydrominer modification to a longwall shearer unit, wherein the rotating pick-laced drums and connecting gear drives are replaced by a water jet cutting ahead. The concept of the water jet cutting head is based on using the high localized energy of the water jets to isolate a slice of coal by cutting a slot under, at the back of, and at the top of the web. The plow blade of the head then advances into the cut and wedges the coal cantilever over onto the conveyor [(Fig. 1)]. There were several potential advantages foreseen for the system; perhaps the most significant were the elimination of coal dust from the working area and the ability, with a variation in nozzle geometry, to cut webs of greater depth with a small increase in haulage horsepower. The University initially designed the head, based on experiments carried out in the laboratory, under an earlier Bureau of Mines contract. The experiments showed that it was more advantageous to work at a pressure of 68.9 mPa (10,000 psi) with an estimated flow rate of 3.15 X 10 - 8 m3/s (50 gpm) than go to the higher pressures and lower flow rates which have been advocated elsewhere.

Jan 1, 1979

By C. M. Marquardt

Tramp iron and steel moving on a conveyor belt cause small currents to be generated in a coil situated in a strong magnetic field, which are converted to an alternating current and are amplified. The output voltage from the amplifier fires a thyratron with a relay in its anode circuit, which actuates a howler and simultaneously drops a spot of marker material on the belt.

Jan 6, 1950

By S. M. Runke, R. G. Meara, O&apos

The Bureau of Mines has been charged by Congress to investigate processes for the production of alumina from low-grade bauxite, alunite, and clay. As one part of the program, an investigation of the application of ore-dressing methods to improve the grade of high-silica bauxite is in progress at the Mississippi Valley Experiment Station at Rolla, Mo. This paper summarizes data obtained so far and presents detailcd results of tests on four typical samples. Mineralogy of Batjxite The name bauxite originally was used to designate the dihydrous aluminum oxide (A1203.2HzO). The designation has been retained, although the existence of a specific mineral of such composition has been disproved. The term is now applied to ores of aluminum consisting of various mixtures of the minerals gibbsite (AlzOa.3H20), diaspore (A120a.H20), and boehmite (Alr-03.H20). The predominant hydrous aluminum oxide in most American bauxite, however, is gibbsite. In some instances bauxite may be amorphous mixtures of the hydrous aluminum oxides that approximate the composition of dihydrous aluminum oxide (A1203.2Hz0). The proposed name for this material is cliachite. Gibbsite and boehmite, the trihydrate and monohydrate, are soluble in a solution of sodium hydroxide at elevated temperature and pressure, whereas diaspore is practically insoluble. The solubility of these two minerals is the basis of the Bayer process of alumina extraction. All bauxites contain certain impurities, the commonest of which, in the order of their importance, are: silica, principally as a constituent of kaolinite, and to a lesser extent as quartz; iron, as hematite, limonite, and siderite; and titanium, as ilmenite, rutile and leucoxene. Minor impurities are zircon, cobalt and nickel. Bauxite is produced commercially in Arkansas, Alabama, Mississippi, Georgia and Tennessee, but the most important deposits in the United States are in Arkansas, in the vicinity of Little Rock. History of Batjxite Beneficiatjon The literature on bauxite beneficiation is limited, but the reports of Gandrud and DeVaneyl and Clemmer, Clemmons, and Stacy2 show the possibilities of flotation as a means of beneficiating bauxite. Their work has also developed reagent combinations that are satisfactory for flotation of gibbsite.

Jan 1, 1944

By W. J. Parton

D. R. MITCHELL*—The Chairman mentioned that we have had many papers on cleaning of fine coal and treatment of wash water solids. There are, of course, two reasons for that. One is that we have legislation, particularly in the east. that is making us go into these things whether we want to or not. It just so happens at the present time that I have four plants that are going up under my general direction, three new and one rehabilitation job, and in all of them, we have these problems. Therefore, I was very happy to be able to listen to these papers, because some of the things that are in them are going to help me solve some of the problems that I am facing. Of course, the second reason why we are concerned about fine coal recovery is the fact that coal has become a luxury item in a good many places, particularly in the anthracite region, so that we cannot afford to have it go to waste like we did in former years. There is a third minor reason, and that is that we are finding how to handle and use the extremely fine sizes that we did not know how to use in former years. Now, I have a couple of questions that I would like to ask. First, what are the maximum solids that you could have in a pulp going to one of these launder screens? I realize that it would be for anthracite. Second, of course the life of these screen surfaces is not very long— with brass or bronze, I believe it is a week, and with stainless steel, a month. Would it be possible to use screen surfaces like the Bixby-Zimmer round rod for the small sizes and possibly increase the life? Have you experimented with that? Another point that I would like to bring out, both in Mr. Parton&apos;s paper and in some of the others, in studying cyclones and any settling apparatus where we use orifices, the life of the orifices is very short, and that is a problem that still has to be solved. The capacity of the cyclone seems to be small, and if you have a bat-

Jan 1, 1950

By J. J. Gilman

Examination showed that characteristic cleavage step patterns are observed on the cleavage surfaces of undeformed, slipped, bent, twinned, compressed, and indented zinc crystals; and the effect of temperature is discussed. Dimples were seen to produce cleavage steps in a treelike pattern in otherwise undeformed crystals. The steps seem to originate when cracks intersect screw dislocations. IT has been known for a long time that the path of fracture in polycrystals may be discontinuous (see Jaffe, Reed, and Mannl for review). Recently, Kies, Sullivan, and Irwin2 have proposed, and given evidence, that crack propagation is discontinuous within individual crystals as well. Other evidence has been given by Low.&apos; When discontinuous cracks within a crystal join together to make a macrocrack, the lamellae between each set of two cracks are torn somewhere, forming small cliffs. These cliffs appear as lines when the cleavage surface is observed microscopically.4,5 The lines have been called vein, tree, and riverlike markings by various authors, and they have sometimes been mistaken for fissures. The descriptive term cleavage steps is used in this paper. Cleavage steps vary in height over a wide range of values, from molecular dimensionsG to lor. and larger. Kies, Sullivan, and Irwin,2 as well as George,&apos; have shown that the gross cleavage step patterns for plastics, polycrystalline metals, and for mono-crystals are sometimes similar. Thus, they depend mostly on the mechanical variables that prevail during cleavage and are relatively insensitive to the structure of the material. For example, parabolic markings2,7,8 sometimes result when cracks open up ahead of, and not coplanar with, the main crack front. If the advance crack has the same velocity as the main crack, their intersection line is a parabola, otherwise it is a hyperbola or an ellipse. The patterns are strongly affected by differences in crack velocities. This results in chevron patterns which point to the place of origin of the main crack. It is the purpose of this paper to demonstrate the existence of a mechanism of cleavage step formation which is a continuous rather than a discontinuous process. Also, certain characteristic step patterns are described, and the strong effect of temperature is shown. The specimens were zinc monocrystals (grown from 99.999+ pct pure metal). These were cleaved at room temperature and at — 196°C. Results and Discussion Cleavage step patterns are highly variable from point to point on a given specimen, as well as from one specimen to another. Although the patterns shown in the photographs are typical, they have been selected for graphic illustration. Figs. la and lb compare undeformed crystals that were cleaved at —196 °C and room temperature, respectively. Cleavage at room temperature (Fig. lb) resulted in a higher density of high steps (dark black lines) and enhanced the visibility of the fine background markings. Deformation by simple slip caused no marked change in the step patterns until the glide strain reached about 1.0. But, as Fig. lc shows, the density of high cleavage steps was greatly increased by large glide strains. Corrugations lying perpendicular to the slip direction may also be seen in Fig. lc. These are caused by deformation bands. The cleavage resistance of the crystal of Fig. lc was very high compared to undeformed crystals (estimated by the force on a needle required for cleavage). Striking and varied cleavage step patterns were observed on bent crystals. Two characteristic patterns that were observed on crystals bent at 25°C, and cleaved by reverse bending at —196°C, are shown in Figs. 2a and 2b. The first, Fig. 2a, consists of V-shaped lines similar to the parabolas of other materials2,7 Fig. 2b shows a pattern that is the equivalent of Fig. la, consisting of faint background lines with a few higher step markings. Cleavage of bent crystals at room temperature resulted in Figs. 2c and 2d. Now, the cleavage step lines show a strong tendency to follow one of two perpendicular paths. In Fig. 2c (bent once), many of the cleavage step components that lie parallel to the bend axis are assembled into irregular lines. In Fig. 2d (bent twice), the cleavage steps again tend to consist of two perpendicular components, but neither of the components is assembled into lines. Also, the step density is higher.

Jan 1, 1956

By Robert Warner

WHEN a new mine is opened, and often when an operating mine must adapt itself to physical or economic changes, a mining system must be selected in complete detail. In the past the plan chosen was usually the product of the local staff alone, however limited the training and experi-ence of that staff might be, which often resulted in the introduction of methods that were familiar to the staff, but not so well suited to the con-ditions as others unknown to the local men. A new management, noting the difficulties with the existing layout, might make modifications or complete changes, so that many mines in their history have seen a whole series of mining systems, each somewhat laboriously grafted on to its predecessor. This should not be interpreted as a condemnation of the older staffs for sticking to the tried systems, such as open square-sets in the early days at the Homestake, because they knew it would work after a fashion, nor does the statement apply to the technically trained, and usually broadly experienced staffs of most present-day properties. The older miners did not have the variety of tools and methods, nor the present freedom of transfer of information as to what the other man is doing, and so did the best they could under the circumstances. Finally, the higher grade of the ores made the choice less vital. The ever-increasing need to mine low-grade orebodies, with a fixed or even lower price for the metal produced, has caused an increasing need for low-cost mining. In most new mines this has simplified the problem of selection by eliminating from consideration all of the high-cost systems, but it has also complicated the choice by the necessity for refinement in detail and a close adjustment of the plan selected to all the conditions at the property. In these low-cost problems any textbook division of mining systems as applied to "beds, veins or masses" avails little as a guide. It should be realized at the outset that to&apos; be successful under modern competitive conditions a system of mining must be worked out with elaborate detail, with careful time scheduling and control of the various elemental operations, and that after it is introduced it must be run with a large amount of technical direction.

Jan 1, 1933

Stripping of the Exotica open-pit mine, which is proceeding at about 120,000 tpd, is dwarfed by its proximity to the 500-million-ton Chuquicamata tailing pile. The excavating work is being carried on smoothly and efficiently and, as of the end of October, accounts for the removal of 77 million tons of total material of which some 200,000 tons is stockpiled ore. Diamond drilling has outlined an orebody 5900 ft long and 3400 ft wide containing an estimated 153 million tons of ore at 1.35% recoverable copper based on 80 % extraction.

Jan 11, 1969