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By N. A. MacLeod

Ball wear in wet grinding mills has been the subject of considerable discussion in the last few years. Its importance to millmen is obvious in view of the fact that ball wear may cost from 2 to 4 per ton of ore for the more fortunate operators and many times this cost when a refractory ore must be reduced to a fine state of comminution. Consequently, any factual information that may be obtained on the nature of ball wear may be applied to introduce changes leading to lower ball consumption and therefore also to lower costs. In 1940, while employed with the Britannia Mining and Smelting Co., Britannia Beach, B.C., I had occasion to investigate factors that might affect ball consumption. We performed a number of tests, which will be briefly described. These tests originated as a result of changing from a 3½-in. ball to a 2-in. ball of different type of manufacture and were motivated by the desire to obtain the best possible type of 2-in. ball. Our testing was helped by the fact that the company operated a one-ton electric furnace, so that it was quite easy to make steel changes. It should be emphasized at the beginning that these tests produced results that in their general application to milling should be interpreted qualitatively rather than quantitatively. As will be evident in the description of the method used in testing, the balls were subjected to an intermittent batch grinding treatment, which may have produced results in rate of steel wear that may not have an accurate correlation with steel consumption in full scale mill operation. However, with this bed and rather arbitrary method of testing, results were obtained that may be of some interest and may stimulate further work on this subject. Unfortunately, owing to pressure of war work, our tests, which we had intended to repeat on a continuous grinding operation, had to be discontinued. Method of Testing Tests Were made in a Small cylindrical mill, rotating at 72 r.P.m. Internal dimensions of the mill were 10½-in. dia. by 9 in. The cylindrical inner surface was lined with rubber and the two ends were formed by steel plates. A charge of 2000 grams dry feed and 2000 C.C. of water was added evev 2 hr., the mill being drained of pulp before the addition of each charge, by tilting and sluicing with water. The mill was run On afternoon and night shift for daily periods of about 15 hr. total grinding time, and each test lasted from 60 to 150 hr. This batch method of testing involved a higher average pulp temperature in the mill, which undoubtedly affected comparative results in Some of the tests. Hardness oF Ball vs. Wear In these tests, special "eutectoid" steel balls were used, some of which were chill-Cut and quenched, some sand-cast and quenched, and some sand-cast and annealed to a dead soft condition. The results obtained, using as mill feed an ore containing considerable free quartz and

Jan 1, 1947

By N. A. MacLeod

Ball wear in wet grinding mills has been the subject of considerable discussion in the last few years. Its importance to millmen is obvious in view of the fact that ball wear may cost from 2 to 4 per ton of ore for the more fortunate operators and many times this cost when a refractory ore must be reduced to a fine state of comminution. Consequently, any factual information that may be obtained on the nature of ball wear may be applied to introduce changes leading to lower ball consumption and therefore also to lower costs. In 1940, while employed with the Britannia Mining and Smelting Co., Britannia Beach, B.C., I had occasion to investigate factors that might affect ball consumption. We performed a number of tests, which will be briefly described. These tests originated as a result of changing from a 3½-in. ball to a 2-in. ball of different type of manufacture and were motivated by the desire to obtain the best possible type of 2-in. ball. Our testing was helped by the fact that the company operated a one-ton electric furnace, so that it was quite easy to make steel changes. It should be emphasized at the beginning that these tests produced results that in their general application to milling should be interpreted qualitatively rather than quantitatively. As will be evident in the description of the method used in testing, the balls were subjected to an intermittent batch grinding treatment, which may have produced results in rate of steel wear that may not have an accurate correlation with steel consumption in full scale mill operation. However, with this bed and rather arbitrary method of testing, results were obtained that may be of some interest and may stimulate further work on this subject. Unfortunately, owing to pressure of war work, our tests, which we had intended to repeat on a continuous grinding operation, had to be discontinued. Method of Testing Tests Were made in a Small cylindrical mill, rotating at 72 r.P.m. Internal dimensions of the mill were 10½-in. dia. by 9 in. The cylindrical inner surface was lined with rubber and the two ends were formed by steel plates. A charge of 2000 grams dry feed and 2000 C.C. of water was added evev 2 hr., the mill being drained of pulp before the addition of each charge, by tilting and sluicing with water. The mill was run On afternoon and night shift for daily periods of about 15 hr. total grinding time, and each test lasted from 60 to 150 hr. This batch method of testing involved a higher average pulp temperature in the mill, which undoubtedly affected comparative results in Some of the tests. Hardness oF Ball vs. Wear In these tests, special "eutectoid" steel balls were used, some of which were chill-Cut and quenched, some sand-cast and quenched, and some sand-cast and annealed to a dead soft condition. The results obtained, using as mill feed an ore containing considerable free quartz and

Jan 1, 1947

By W. K. Hardy, H. H. Stadelmaier, L. J. Huetter

The ternary system Co-Sn-C was studied in the cobalt-rich region. The liquidus projection ulas established and two fonr -phase reactions were found. An isothermal section at 950°C was worked out. In the region studied the following phases were observed: cobalt-rich solid solution, graphite, binary compound y -cobalt-tin, and a ternary compound. As a result of the limited solubility of carbon in liquid cobalt-tin alloys, the investigation was limited to the cobalt-rich region. Tin contents up to 32 at. pct were studied. Cobalt of 99.99 pct purity, tin of 99.97 pct purity and synthetic graphite of 99.93 pct minimum purity prepare the alloys. Carbon determinations of all alloys were obtained by gas volumetric analysis. Samples were investigated microscopically and by X-ray diffraction using CoK, and FeK, radiation. Cooling curves were obtained with a number of alloys to determine the temperature of the four-phase reactions. The thermocouple was calibrated with the melting points of high-purity silver and copper. In the region studied the following phases were observed: Cobalt-rich solid solution (face-centered cubic) Graphite. Binary compound y (solidifies congruently at composition Cop~n)'. Ternary compound k which is face-centered cubic with superlattice and nonferromagnetic.' In Fig. 1 the fields representing the projections of the liquidus surfaces were determined by microscopic investigation and are labled by their primary phase. In the field labled C graphite is the primary phase. During solidification in this field much of the primary graphite was expelled through the surface of the melt. It could not always be established microscopically whether the remaining graphite was primary or purely eutectic. Therefore it was not possible to determine accurately the boundary curve which drops from the binary eutectic el at 1309 "~~ to the point of the four-phase reaction at F. The ter-nary carbide is the primary phase in the field labled k. To the right of this field the primary graphite is consistently nodular whereas flake graphite is observed at tin contents somewhat below F. The ternary carbide does not appear to solidify congruently. The graphite is attacked in a peritectic reaction along the boundary between k and C. Thus no graphite is observed in alloys in the k field. In alloys right of the boundary between k and C the graphite nodules are located in the center of the k crystallites, further documenting the peritectic nature of the reaction with the graphite. A four-phase reaction was found at F and 1120 &3"C. The most likely reaction is L+ CY-Co + C — k. Below the temperature of the binary eutectic ez at 1112" a second four-phase reaction was observed at E and 1100 lt3"C. It is the ternary eutectic L — CY-Co + k + y. Figs. 2(a ) to (e) show microstructures of the ternary alloys in the as-cast condition. Fig. 2(a) shows the eutectic a-Co + C (with some primary graphite) with the carbide k in the grain boundaries. Fig. 2(b) shows primary cobalt (dark areas) and the eutectic a-Co + k which is formed along F-E.

Jan 1, 1961

By H. Buttner, F. H. Udin, J. Wulff

GRAIN boundaries, the interface between adjacent crystals differing only in respective orientation, have been the object of much experimental and moderate theoretical attention for many years. The earlier experimental work was primarily concerned with the effect that grain boundaries have on the physical behavior of metals, while the basic nature of the boundaries themselves drew only secondary attention. A complete turnabout has been witnessed in more recent years, however, as researchers are now undertaking investigations bearing as directly as possible on the character and physical properties of the grain boundary. Quite comprehensive reviews of the past literature on the subject of grain boundaries have been reported on numerous occasions by various authors, and together these references represent a thorough coverage.'-? It becomes evident in these reviews and other papers"" that the properties of grain boundaries can be explained by regarding a grain boundary as a crystalline region of transition between two unmatched grains. Studies show that grain boundaries have a definite interfacial energy. By measuring the geometry about a point where three interfaces meet in equilibrium, the energy of any two boundaries can be measured with respect to the third. If the absolute interfacial energy of one boundary is known, the absolute energy of the other two boundaries can be calculated. This apparently holds true for any system of phases and interfaces.""25 Most studies have been carried out on solid-solid or liquid-solid systems, or systems of both. However, if about the point of intersection one phase is vapor and the other two are disoriented grains, then two interfaces are surfaces, and one is a grain boundary. Investigations leading to relative grain boundary energy for this particular system have been made by Fullman,23 Greenough and King," and Bailey and Watkins." Herring" treats the system mathematically and arrives at an equation, which appears as Eq. 1, in a form consistent with the reference angles indicated in Fig. 1, ( d?1 ?n = x sina + cosa ( ------- 2 ?2 Sinß da + C0Sß (d?2/d As the system of interfaces comes to equilibrium, a groove of angle appears at the intersection of the grain boundary with the surface, an effect first observed by Rosenhain and Humfrey." Eq. 1 is essentially a summation of the vertical components in Fig. 1 set equal to zero, where ?1 and ?2 are surface tension vectors lying along the side of the groove at point 0 in the groove root. The differential vectors represent the forces tending to rotate the surface into a position that will expose a crystal cleavage surface of lowest energy. Chalmers, King, and Shuttleworth- simplify Eq. 1 by assuming the differential vectors equal to zero, ?1 = ?2= ?Surface and arrive at the following equation: ?B = 2?s cos [2] Greenough and King24 use Eq. 2 for determining the grain boundary energy of silver using prepared bi-crystals of known orientation. Bailey and Watkins25 have done the same for copper, as did Chalmers and others on lead. These authors report the grain boundary values as ratios relative to the surface energy. In the above work, the orientation of grains forming the grain boundary in all cases was controlled so

Jan 1, 1954

By John M. Bostwick, James C. Bradbury

Bitumen is a general term for a group of materials composed of mixtures of hydrocarbons that are fusible and soluble in carbon disulfide. Included in the group are petroleum, asphalts, asphaltites, and mineral waxes. Asphalt is a dark-colored, cementitious solid or semisolid, which generally softens on heating and resumes a more viscous or solid state when cooled. It is found in native form in fissures and pore spaces in rocks and as lakes. It is also obtained from the residues produced in petroleum refining. Asphaltites are black, naturally occurring solid bitumens. They are generally shiny and break with a conchoidal fracture. Common asphaltites are gilsonite (uintahite), grahamite, and glance pitch. Mineral waxes are solid or semisolid, relatively nonvolatile bitumens of various colors, having an unctuous feel and waxy luster, Of the native waxes, ozokerite is the best known. Montan wax is also classed as a natural wax, although it never occurs in the pure state and must be extracted from lignite or cannel coal (Ladoo, 1920). Mineral waxes may also be obtained by pyroprocessing of petroleum, lignite, and peat. Pyrobitumens are dark-colored, solid hydrocarbons that are infusible and relatively insoluble in carbon disulfide. They are so-called because they will yield liquid and gaseous hydrocarbons only when heated. Asphaltic pyrobitumens are defined as "substantially free of oxygenated bodies" and include wurtzilite (elaterite), albertite, and impsonite (Thrush, 1968). Nonasphaltic pyrobitumens contain oxygenated bodies; examples are coal and peat. The nonasphaltic type will not be discussed further. Petroleum Asphalts Asphalts refined from crude petroleum have almost completely replaced the native types. Refineries can produce asphalts in every desirable grade more economically and in greater quantity than the native types and with a variety of properties that meet the requirements of the various manufacturers using them. For all practical purposes, it is estimated that more than 95% of the asphalts consumed in the United States are refined from crude petroleum (private communication). Native Asphalts Native asphalts are dark brown or black, semisolid or solid substances that, in theory, formed by slow, natural fractionation of crude petroleum at or near the surface of the earth (Goulston, 1938). They are colloidal systems of high molecular weight components dispersed in a medium consisting of low molecular weight components (Pracy, 1954). Native asphalts may occur as lakes, as impregnations of rock (rock asphalt) or as deposits consisting of mixtures of natural asphalt and mineral matter in various proportions (Stern, 1960). Practical use of native asphalt dates to about 3000 B.C. It was mined from surface pits or lakes and used, then as now, for masonry mortar, waterproofing aqueducts and tunnels, binders for road construction, and as an article of commerce (Miller, 1938). Generally the native asphalts find little commercial use today except where they are geographically abundant, easily mined and transported, or have specific or special application. Petroleum-refined asphalt is produced by unusually economical techniques and is much less costly to use. The largest and best known of the lake deposits is that on the Island of Trinidad, where the lake covers an area of 114 acres and is 285 ft deep at its center (Stem, 1960). The lake asphalt is refined by heating to evaporate in-

Jan 1, 1983

By F. Weinberg

The distribution of solute during the progressive solidification of dilute Sn-Ag alloys was determined in both solid and liquid as a jbnction of growth rate, rod diameter, temperature gradient, and solute concentration, using radioactive tracer techniques. The results indicated that partial mixing occurs in the liquid during freezing for most of the conditions investigated, approaching complete mixing for larger rod diameters (>2mm) and slow growth rates. Measured solute distributions were fitted to distribution functions based on diffusion controlled solute transport in the liquid state, by treating the normal solute diffusion constant and distribution coefficient as variable solute transport and segregation parameters. Values of these parameters were determined for the growth conditions investigated. Observations were made to determine whether mixing in the liquid was primarily due to volume changes in freezing or to differences in density between solute and solvent. The former effect was investigated using a selected Sn-Bi alloy as a solvent with zero volume change on fusion, and the density effect was studied using vertical solidifications of Sn-In and Sn-Ag alloys, systems having, respectively, very small and large differences in density between solute and solvent. The result indicate that mixing in the liquid cannot be primarily attributed to either volume changes or density differences. THE solute distribution in progressively solidified rods of dilute alloys, in which there is complete mixing in the liquid, is given by the expression1 where C, is the solute concentration in the solid at the point where the fraction g of the material has solidified, C, is the initial solute concentration and k is the equilibrium distribution coefficient. If it is assumed that there is no significant con-vective mixing in the liquid, then the solute distribution in the initial portion of a semi-infinite rod has been calculated to be2 where t is the time from the start of freezing and x1 is the distance in the liquid ahead of the solid-liquid interface at time t. Expressions [1] and [2] describe the limiting cases of solute distribution in the liquid, i.e., complete mixing, or no convective mixing; in practice it might be expected that the solute distribution would be somewhere between these two cases. pfann3 has suggested that Eq. [I] is applicable for partial mixing, as well as complete mixing, if an effective distribution coefficient k1 is used instead satisfactorily describes solute distributions in rods of magnesium, n-g, and b-sb6 alloys that were progressively solidified from the melt without mechanical mixing. The values of kl were observed to depend on the growth rate4-6 and the temperature gradient.6 The purpose of the present investigation was to measure solute distributions in both the solidified rods and the melt as a function of the solidification variables, and to determine to what extent the observed solute distributions could be described by the above theoretical expressions, keeping in mind that expressions [2] and [3] have been derived solely for the case of diffusion controlled solute transport in the liquid and are not theoretically applicable to

Jan 1, 1963

By Michael P. Richardson

INTRODUCTION This paper is intended to be a summation of the state-of-the-art i n draw theory and draw- point spacing. Knowledge of this very important factor in the successful design and operation of a block caving mine is scarce and nonconclusive. It is hoped that the presentation of this paper will stimulate interest and discussion which will lead to a stronger understanding of these a1 1 important factors. Please note that block caving as used here- in refers to block, panel, and mass caving systems in general, not just to the specific system of block caving where discrete blocks of ore are prepared and caved as a unit. DRAW THEORY The knowledge of draw theory as related to the practice of block caving mines is, unfortunately, very limited. In recent years much work in draw theory has been done in respect to sublevel caving systems and an understanding of draw mechanics relating to sublevel caving seems to be emerging. This work is certainly helpful to those of us interested in block caving but the question of "what goes on up above" still hounds us since we are more interested in the region of draw which begins just above the area where sublevel caving interest ends. Herein is presented a brief summary of draw theory with an attempt to look at the factors which can help our understanding of the area of draw influence in block caving mines. Nature of Gravity Flow An understanding of the flow of granular material (rock, sand, grain, or prill as examples) is the first step toward understanding block caving draw. If we take a supported column of granular material and allow gravity flow through an aperture at the bottom of the column, an ellipsoidal shaped zone of disturbance of the in-place material will form and grow with extraction of material through the aperture. The volume of this zone of disturbance or loosening is about 15 times the volume of material extracted at any given time (Kvapil, 1980). Figure 1 shows the development over time of the zone of disturbance. The ellipsoid of disturbance represents a future ellipsoid of extraction which will equate to the volume of materia1 extracted by that future time. Extraction involves the movement of the particles located vertically over the drawhole directly downward while material outside this direct vertical volume moves radially inward and down toward the drawpoint as shown in Figure 2a. Figure 2b shows the result of this motion on originally horizontal strata within the ore column. Increased draw increases the eccentricity of the ellipsoidal volume or, in other words, the ratio of the height over the width of the disturbed area increases as ore is withdrawn from the column (Kvapil , 1980). This becomes important to block caving since eventually a vertical cylinder of draw will be approximated as this ratio increases without bounds. Figure 3 shows this result. The width of this draw cylinder is a critical dimension for the determination of draw point spacing in relationship to block caving mine design. McCormick (1968) for one, felt that within this cylinder of draw, mass flow would take place. This is the

Jan 1, 1981

By R. L. Robinson, P. DeFrank, R. F. Hatfield

Production equal to or greater than open-hole completions is possible through perforated completions if the flow paths throughout the perforations are free of obstructions.' Previous investigations have indicated that, in some cases, maximum productivity is not obtained due to plugging of the perforations with debris from the wellbore fluid and/or debris from the perforating process. Control of wellbore fluid and proper well completion techniques are very important in the elimination of perforation plugging.' Selection of properly designed perforating equipment also is quite important for maximum well productivity. Studies made with the use of the perforation flow laboratory, in which perforating tests are run under simulated well conditions, have shown that perforation plugging occasionally occurs. Plugs in perforations result from various causes. One of the contributors to plugging that can be controlled by the perforating device is the slug or carrot resulting from jet perforating. This condition has undergone extensive study, and previous developments have resulted in a decrease in carrot size—but not in complete elimination. It was felt that, if a method of completely eliminating the carrot could be found (with charge performance equal or better than previously obtained), a significant contribution would have been made. Such a charge has been developed and is reviewed in this paper. THEORETICAL CONSIDERATIONS OF CARROT ELIMINATION There are several popular theories of how a shaped charge jet is formed."," One of these is the hydrodynamic theory, wherein the liner is treated as a fluid under the impact of the explosive forces. Another is the spalling theory, where the interaction of shock waves in the liner causes small particles to be torn from the inner surface of the liner and projected at high velocity toward a common focal line to form the jet. It is not unreasonable to assume that both of these theories may be partially correct. For purposes of our discussion, it is simpler to use the second theory. Generally speaking, when a plate of metal is attacked by an explosive shock pulse, the pulse is transmitted through the metal at high velocity, as shown in Fig. 1(A). Pressures at the shock front may be 6 X 10' psi or higher. This shock wave (a compression wave) upon striking a free surface, Fig. 1(B), is partially reflected back as a tension wave. The interaction of the tension wave with the decaying compression wave causes a stress concentration in the metal; when this stress becomes great enough, a failure occurs and a small particle flies off and is projected away from the metal surface, as shown in Fig. 1(C). A new free surface is then presented to the shocks; the process repeats itself until there is no longer enough energy to cause failure, leaving a portion of the metal relativcly untouched. See Fig. 1(D). To apply this theory to a perforating shaped charge, consider the flat plate to be folded into a somewhat conical shape, with a column of explosive placed in intimate contact with the cone (Fig. 2). An explosive initiator detonates the charge, creating a shock wave

By R. M. Hardgrove

Pulverized coal was first used for firing cement kilns about 45 years ago, with such success that it has continued in general use. Based on cost, pulverized coal is usually the most economical fuel. In some localities natural gas or oil is cheaper, but 85.5 per cent of all cement kilns at present are fired with pulverized coal. The pulverized-coal flame is well suited to kiln conditions. Its soft, luminous nature gives it a high radiating property and the length of the flame can be readily varied to suit conditions in a kiln of any length. Ash and slag cause most of the dificulties incident to the burning of coal. In a. cement kiln, however, the ash, being similar in composition to the raw mix and being deposited uniformly in the mix and clinker, occasions no difficulty as long as it remains constant in quality and quantity. Until 6 years ago almost all cement kilns using pulverized coal employed the bin system. The coal was dried and pulverizeti in a separate plant and delivered into a bin at the kiln. Thus the bin system requires driers, storage-type pulverizers, conveying systems, individual kiln bins, pulverized-coal feeders and primary air fans. The direct-fired system requires only a pulverizer and a blower. Fig. I shows typical plants. The first installations of direct-fired pulverizers were of the impact type but fineness of pulverization was both inadequate and variable. About 1927 a few direct-fired installations employing tube-type pulverizers were made, which were much more satisfactory. As more efficient pulverizers were developed from 1934 on, equipment for direct firing was installed at a large number of plants, until now 196 of about 900 kilns in use in the United States are direct-fired. Most of these replaced antiquated bin systems. Process of Burning Cement Clinkers The process of burning cement clinkers is fundamentally one of heat transfer. In boilers and other heat-exchange apparatus the area of the surface receiving heat from the gases is the best unit for comparing output, This seems to be equally true for kilns. The inside kiln surface is the heat-receiving surface and its area is used as the unit of kiln size in the following comparisons. Heat input is based on British thermal units in the fuel. E. G. Bailey,' in his work on boiler-furnace heat absorption, has used the unit "B.t.u. available." This is the practical calorific power of the fuel, plus the heat in the air used for combustion. In other words, it is the heat available for useful work in the furnace or kiln. The unit of kiln output is barrels per hour. The composition of the raw mix, the percentage of water in the slurry of wet plants and hardness of burning all affect the output. These conditions, however, are hard to evaluate and probably have less effect on the output than irregularities of raw feed and clinker discharge and inexpert control of combustion. Wet vs. Dry Plants The study presented here covers 88 plantsl of which 41 are wet and 47 dry. Of

Jan 1, 1942

By R. M. Hardgrove

Pulverized coal was first used for firing cement kilns about 45 years ago, with such success that it has continued in general use. Based on cost, pulverized coal is usually the most economical fuel. In some localities natural gas or oil is cheaper, but 85.5 per cent of all cement kilns at present are fired with pulverized coal. The pulverized-coal flame is well suited to kiln conditions. Its soft, luminous nature gives it a high radiating property and the length of the flame can be readily varied to suit conditions in a kiln of any length. Ash and slag cause most of the dificulties incident to the burning of coal. In a. cement kiln, however, the ash, being similar in composition to the raw mix and being deposited uniformly in the mix and clinker, occasions no difficulty as long as it remains constant in quality and quantity. Until 6 years ago almost all cement kilns using pulverized coal employed the bin system. The coal was dried and pulverizeti in a separate plant and delivered into a bin at the kiln. Thus the bin system requires driers, storage-type pulverizers, conveying systems, individual kiln bins, pulverized-coal feeders and primary air fans. The direct-fired system requires only a pulverizer and a blower. Fig. I shows typical plants. The first installations of direct-fired pulverizers were of the impact type but fineness of pulverization was both inadequate and variable. About 1927 a few direct-fired installations employing tube-type pulverizers were made, which were much more satisfactory. As more efficient pulverizers were developed from 1934 on, equipment for direct firing was installed at a large number of plants, until now 196 of about 900 kilns in use in the United States are direct-fired. Most of these replaced antiquated bin systems. Process of Burning Cement Clinkers The process of burning cement clinkers is fundamentally one of heat transfer. In boilers and other heat-exchange apparatus the area of the surface receiving heat from the gases is the best unit for comparing output, This seems to be equally true for kilns. The inside kiln surface is the heat-receiving surface and its area is used as the unit of kiln size in the following comparisons. Heat input is based on British thermal units in the fuel. E. G. Bailey,' in his work on boiler-furnace heat absorption, has used the unit "B.t.u. available." This is the practical calorific power of the fuel, plus the heat in the air used for combustion. In other words, it is the heat available for useful work in the furnace or kiln. The unit of kiln output is barrels per hour. The composition of the raw mix, the percentage of water in the slurry of wet plants and hardness of burning all affect the output. These conditions, however, are hard to evaluate and probably have less effect on the output than irregularities of raw feed and clinker discharge and inexpert control of combustion. Wet vs. Dry Plants The study presented here covers 88 plantsl of which 41 are wet and 47 dry. Of

Jan 1, 1942

By J. P. Givry

The bath and the liquid metal in the aluw~inuw~ electrolytic cells are moving under the action of the Laplace forces. Consequently there are differences in level at the interface of the liquid. This phenomena is studied by means of a computer; the characteristics of the mathematical model used are presented with the results corresponding to two typical cases. This study gives a new and origi~zal interpretation of magnetic ej-jects; their action depends very closely on the hydraulic phenomenon which explain the sensibility of the cells to the slope and thickness of solid bath frozen around the pot. MAGNETIC effects produced in the electrolytic cells used for the production of aluminum are known to exist, a factor whose importance is well-known by producers through its influence on cell operation and output. In particular, these effects bring about movements in both the bath and metal as well as interface variations in level that are detrimental to sound electrolysis . A direct study of these effects by means of an experimental device would raise a number of problems particularly difficult not to say impossible to solve in the present state of techniques. It is for this reason that we had recourse to an electronic computer for working out the speed and pressure distribution within the liquid in the cells. We have assumed that all other possible causes for movement, such as thermal convection and evolution of gas at the anode, are acting without any connection with magnetic effects, and that consequently we may study each of these causes separately. A basic cause for magnetic effects lies in the presence of Laplace forces within the bath and metal. Electric current and magnetic field distribution have been calculated with the help of an electronic computer, and two communications about it have already been made to AIME.* 1) PRESENTATION OF THE MATHEMATICAL PATTERN In order to solve this problem, it was essential to assume a number of simplifying hypotheses with regard to both the geometry of the actual cells and the phenomena themselves. A first simplification consisted in assuming that the movements of the conducting liquid modify neither the electric currents nor the magnetic fields. In consideration of the magnitude of the values involved this assumption seemed accep-tible. 1.1) Finding the Equation of the Problem. The problem raised consisted in determining the flow of two viscous immiscible liquids with different densities subjected in a complex environment to gravity and Laplace forces. As a matter of fact, the sum of the external forces acting on the liquids is: F = JAB +pg where F = volume energy, B = magnetic field, J = current density, p = specific gravity, g = gravity. In the case of a laminar flow and assuming that the liquids are Newtonian and incompressible liquids, the equations governing movements are the following: 1) Navier Stokes' equation written in vectorial form: In fact one tries to find a steady-state solution, so aVx/at = 0; besides, one assumes that the movements are slow enough for quadratic terms such as Vy(a Vx/ay) (of round about the square of speed) to be neglected in front of the viscosity term: pAV, so

Jan 1, 1968

By W. C. Knoblaugh

Rotary kilns are adaptable to many fuels, but this paper deals principally with the use of powdered coal. The observations and conclusions presented are based on rotary kilns used in the manufacture of basic refractories for steel furnaces. Solid fuels are pulverized for two principal reasons: (I) ease of control, and (2) rapidity of combustion. By the use of pulverized coal, the operator can increase or decrease the fuel supply by small increments and with very little time lag between the change of control and the resultant change in the flame. This is extremely important in this type of furnace, as the temperature range for maturing the clinker is very narrow. By fine grinding, it is possible to increase the rate of burning of the coal, thereby liberating great quantities of heat within a small space in a short time. Under given conditions, temperature is a measure of the heat liberated per unit of time. To attain the temperatures of 2950° to 3150ºF., which are encountered in the manufacture of various steel-furnace refractories, it is estimated that it is necessary to liberate 125,000 to 175,000 B.t.u. per cubic foot of the burning zone per hour. The manufacture of dolomitic refractories is a ceramic process and, as is true in most such processes, the material must be subjected to a certain critical temperature for a certain length of time in order to complete the reaction. If the time of reaction is reduced, the temperature must be increased; in other words, if the rate at which the material flows through the hot zone of the kiln is increased, the temperature of the hot zone must be increased to compensate for this shorter exposure to the flame. Firing with Pulverized Coal Fig. I is a line drawing of a rotary kiln with the attendant equipment indicated in their relative positions. The kiln proper consists of a sloping, refractory-lined steel tube rotating slowly about its longitudinal axis (0.75 r.p.m.), and is countercurrent in operation in that the raw feed is introduced at the stack end and gradually approaches the zone of highest temperature at the lower end, where the fuel is introduced. Before 1930, the usual method of firing a rotary kiln was to store the coal in the pulverized form and then feed from this storage to the kiln as needed. This proved unsatisfactory and dangerous because of dust explosions and the inability to obtain an even feed to the kiln caused by arching of the coal in storage and flooding of the coal feeders. In later years the practice has been to store the raw coal and grind it as it is needed. Such a layout is indicated in Fig. I. The coal is stored in bins, charged from the top and drawn from the bottom into recording scales and then into some type of unit pulverizer. There are a number of good pulverizers .on the market, each incorporating features that the manufacturer considers outstanding.

Jan 1, 1948

By W. C. Knoblaugh

ROTARY kilns are adaptable to many fuels, but this paper deals principally with the use of powdered coal. The observations and conclusions presented are based on rotary kilns used in the manufacture of basic refractories for steel furnaces. Solid fuels are pulverized for two principal reasons: (I) ease of control, and (2) rapidity of combustion. By the use of pulverized coal, the operator can increase or decrease the fuel supply by small increments and with very little time lag between the change of control and the resultant change in the flame. This is extremely important in this type of furnace, as the temperature range for maturing the clinker is very narrow. By fine grinding, it is possible to increase the rate of burning of the coal, thereby liberating great quantities of heat within a small space in a short time. Under given conditions, temperature is a measure of the heat liberated per unit of time. To attain the temperatures of 2950° to 3150°F., which are encountered in the manufacture of various steel-furnace refractories, it is estimated that it is necessary to liberate 125,000 to I75,000 B.t.u. per cubic foot of the burning zone per hour. The manufacture of dolomitic refractories is a ceramic process and, as is true in most such processes, the material must be subjected to a certain critical temperature for a certain length of time in order to complete the reaction. If the time of reaction is reduced, the temperature must be increased; in other words, if the rate at which the material flows through the hot zone of the kiln is increased, the temperature of the hot zone must be increased to compensate for this shorter exposure to the flame. FIRING WITH PULVERIZED COAL Fig. I is a line drawing of a rotary kiln with the attendant equipment indicated in their relative positions. The kiln proper consists of a sloping, refractory-lined steel tube rotating slowly about its longitudinal axis (0.75 r.p.m.), and is countercurrent in operation in that the raw feed is introduced at the stack end and gradually approaches the zone of highest temperature at the lower end, where the fuel is introduced. Before 1930, the usual method of firing a rotary kiln was to store the coal in the pulverized form and then feed from this storage to the kiln as needed. This proved unsatisfactory and dangerous because of dust explosions and the inability to obtain an even feed to the kiln caused by arching of the coal in storage and flooding of the coal feeders. In later years the practice has been to store the raw coal and grind it as it is needed. Such a layout is indicated in Fig. I. The coal is stored in bins, charged from the top and drawn from the bottom into recording scales and then into some type of unit pulverizer. There are a number of good pulverizers on the market, each incorporating features that the manufacturer considers outstanding.

Jan 1, 1946

By W. C. Knoblaugh

Rotary kilns are adaptable to many fuels, but this paper deals principally with the use of powdered coal. The observations and conclusions presented are based on rotary kilns used in the manufacture of basic refractories for steel furnaces. Solid fuels are pulverized for two principal reasons: (I) ease of control, and (2) rapidity of combustion. By the use of pulverized coal, the operator can increase or decrease the fuel supply by small increments and with very little time lag between the change of control and the resultant change in the flame. This is extremely important in this type of furnace, as the temperature range for maturing the clinker is very narrow. By fine grinding, it is possible to increase the rate of burning of the coal, thereby liberating great quantities of heat within a small space in a short time. Under given conditions, temperature is a measure of the heat liberated per unit of time. To attain the temperatures of 2950° to 3150ºF., which are encountered in the manufacture of various steel-furnace refractories, it is estimated that it is necessary to liberate 125,000 to 175,000 B.t.u. per cubic foot of the burning zone per hour. The manufacture of dolomitic refractories is a ceramic process and, as is true in most such processes, the material must be subjected to a certain critical temperature for a certain length of time in order to complete the reaction. If the time of reaction is reduced, the temperature must be increased; in other words, if the rate at which the material flows through the hot zone of the kiln is increased, the temperature of the hot zone must be increased to compensate for this shorter exposure to the flame. Firing with Pulverized Coal Fig. I is a line drawing of a rotary kiln with the attendant equipment indicated in their relative positions. The kiln proper consists of a sloping, refractory-lined steel tube rotating slowly about its longitudinal axis (0.75 r.p.m.), and is countercurrent in operation in that the raw feed is introduced at the stack end and gradually approaches the zone of highest temperature at the lower end, where the fuel is introduced. Before 1930, the usual method of firing a rotary kiln was to store the coal in the pulverized form and then feed from this storage to the kiln as needed. This proved unsatisfactory and dangerous because of dust explosions and the inability to obtain an even feed to the kiln caused by arching of the coal in storage and flooding of the coal feeders. In later years the practice has been to store the raw coal and grind it as it is needed. Such a layout is indicated in Fig. I. The coal is stored in bins, charged from the top and drawn from the bottom into recording scales and then into some type of unit pulverizer. There are a number of good pulverizers .on the market, each incorporating features that the manufacturer considers outstanding.

Jan 1, 1948

By G. D. Hobson, I. Mrosovsky

Material balance relationships for expansion-type reservoirs above the bubble point have been discussed in recent years by H. N. Hall and M. F. Hawkins, Jr. The former drew attention to the effects of rock compressibility, and the latter to the effects of the compressibility of any interstitial water. Hawkins derived a general relationship which took note of oil, water and rock compressibilities, and then by assuming first, the absence of interstitial water and secondly, no effects due to interstitial water or rock compressibility, obtained relationships symbolically of the same form as those previously derived directly by Hall. Hawkins defined his terms clearly, and noted an approximation, due to the omission of a term involving the product of two compressibilities, made in reaching his final equation. There is, however, possible ambiguity with regard to Hall's written definition of the oil production term. One interpretation would mean that the produced oil and initial oil-in-place, if both required as stock tank barrels, would be obtained by using different formation volume factors in the conversion in order to satisfy the conditions implicit in Hall's formulae. The other would involve an approximation which, if used, was not stated. In both Hall's and Hawkins' studies the effect of the approximation or inconsistency is very slight. Nevertheless, it may be justifiable to give a somewhat more consistent derivation, since the complexity of the final relationships is not thereby greatly increased. Let the initial pore volume be Vi and the volume after a pressure drop of Ap, V. Then V = V, (1 - ?p . cr), where c, is the rock compressibility expressed as change in pore volume/unit pore volume/unit pressure change, and is defined with the initial pressure as base. Suppose that the interstitial water saturation under the initial conditions is Sw. The total volume of interstitial water under those conditions will be ViSw. If the volume of the water after dropping the pressure by ?p is V., this will be given by Vw = ViSw (1 + ?p.c,w), the compressibility of the water being c, which is ex- pressed using the volume at the initial pressure as base. If Boi is the reservoir volume factor of the crude oil for the initial reservoir conditions, and B. the equivalent factor when the pressure has dropped by ?p, with N the initial amount of oil expressed as stock tank barrels. N.Boi = Vi - V,Sw If n is the amount of oil produced (in terms of stock tank barrels) as a result of the pressure drop ?p, the remaining subsurface volume of oil at the lower pressure will be Bw, and Bc can be related to the compressibi:ity of oil and the pressure drop, and if c, is the compressibility factor based on the initial conditions, The relationship derived by Hawkins was case to a standard at the lower pressure, while a term (Ap) had been dropped, because of its smallness arising from its dependence on the products of two compressibility factors. When there is no interstitial water and rock compressibility can be ignored, Eq. 1 reduces to:

Jan 1, 1957

By Nicholas J. Grant, Italo S. Servi

SPECIAL polishing and etching techniques were developed during an investigation of the high temperature mechanical properties of high purity aluminum (99.995 pct Al).* Electrolytic polishing was done successfully in the following electrolyte: 5 ml perchloric acid (72 pct), and 95 ml glacial acetic acid. No special precautions need be taken during the preparation of the electrolyte since no appreciable heat of mixing is evolved. Although the electrolyte is not an explosive, it is recommended that Jacquet's safety rules' be followed during the polishing operation. A voltage of 25 to 60 v gives a good polish in about 2 min at room temperature after preparation on 3/0 metallographic paper. Most of the grain boundaries are visible under the microscope after polishing. Since the current is controlled essentially by the resistance of the bath, it is necessary to keep the anode-to-cathode distance as constant as possible. Accordingly, polishing of irregularly shaped pieces will be difficult. A cooling bath and a stirring system are advisable in order to keep the temperature of the bath uniform and constant. Coarse-grained samples can be macroetched in a few seconds by lowering the voltage below 10 v. This is done after electrolytic polishing, without removing the sample from the bath, Fig. 1. The specimen can be repolished in less than 1 min by raising the voltage above 25 v. The "galvanic etching" technique, which has been described elsewhere,' was found to be helpful not only for macroetching, but also for microetching coarse-grained specimens. In fact, the different orientations of adjacent grains provide good contrast at the grain boundaries, when the sample is observed under polarized light. If the specimen is only partially etched, Fig. 2a, the grain boundaries can be located without completely destroying the appearance of the surface of the specimen. The unetched surface appears black under polarized light. Excellent color contrast is obtained by using a "sensitive tint" plate in the optical system of the microscope. In this case, the unetched surface has a magenta color, while the etched part of the sample assumes different colors varying from green to brown, according to the orientation of the grains. Etch pits were produced on high purity aluminum by using the technique suggested by Lacombe and Beaujard." Good results also were obtained by substituting 70 pct nitric acid for the fuming nitric acid recommended. This gave an etching reagent of the following composition: 70 ml nitric acid (70 pct), 50 ml hydrochloric acid (37.5 pct), and 3 ml hydrofluoric acid (48 pct). An example of the results is shown in Fig- 2b. The outline of the etch pits is revealed sharply if the specimen is observed under polarized light, as shown in Fig. 2c. These metallographic techniques are not suitable for the examination of commercially pure (2s) aluminum. References 'P. A. Jacquet: Metal Finishing (Nov. 1949) 47, p. 62. 21. S. Servi: Metal Progress (1950) 58, p. 732. 3 P. Lacombe and L. Beaujard: Journal Inst. of Metals (1947) 74, p. 1._________________________

Jan 1, 1952

By D. C. Hilty

C. E. SIMS*—This is a most interesting and important paper. It is important from two standpoints. First, it has as-spects of being highly accurate and therefore extremely useful to the operating man in showing how far he can go in oxidizing carbon from a steel bath while retaining chromium and the effect of temperature on this limit. Fig 5 is undoubtedly the highlight of the paper, and in checking it against available data, it seems capable of predicting actual results. The other important aspect is the apparently significant fact that reactions within a steel bath, such as between carbon and an oxide, are homogeneous reactions which do not require equilibrium with an external phase. In this case, it seems obvious that the reaction and the equilibrium established were between dissolved carbon (or carbide) and dissolved CrO. The classic reaction which requires equilibrium with solid Cr2O3, did not fit the data. It was recognition of this principle that made Fig 5 possible. G. R. FITTERER*—It might be in assuming the CrO is dissolved in the metal that you are actually approaching the correct factor for activity. If that is true and if the percentages that were used represent an equilibrium constant, this would mean, in effect, that the minimum temperature at which chromium would be reduced by an equivalent percentage of carbon would be about 2400°F from your equation. I believe that is in line with my experience particularly if you deal only with low percentages. J. CHIPMAN†—The author in discussing the equilibrium of the reaction offers two choices of explanation. One postulates an effect of chromium on the activity coefficient of carbon; it would require perhaps more data than are available to us at the present time to test this thoroughly. The other postulate which he offers is a reaction with CrO. This explanation is consistent with physical and chemical principles provided that a solid phase CrO is present, and that is essentially what he is postulating. If that oxide were present one would expect and, I think, almost require the kind of results that he obtained. The weakness of that postulate lies in the lack of confirmatory evidence regarding the presence of solid CrO as such. This should offer an interesting subject for further study. It seems a pity that the theoretical arguments revolving around CrO should have obscured for a moment the very real excellence of this paper. E. CARTER*—I think there is a problem that presents itself in practical application where the metal temperature is from 3200 to 3300 and the temperature desired in pouring is considerably less. B. M. LARSEN†—I get the impression from some of the work that was done earlier that there is probably a localized zone of high temperature at the slag metal interface which effects the removal of the carbon. Without that I wonder if the effect is possible to have a localized superheat and an equilibrium there favorable to low carbon content and

Jan 1, 1950

By Raymond Mancha

When calculating the pressure-volume characteristics of projected mine-ventilating circuits by orthodox methods, certain basic assumptions are required in order to employ the various available empirical data. It is assumed, for example, that the mine air is an incompressible fluid subjected to isothermal flow, an assumption sufficiently accurate for practical purposes since pressure and temperature differentials are small throughout the average circuit. A more erroneous assumption, however, is that all air is accounted for as it travels throughout the mine. No attempt is usually made to evaluate stopping leakage as it occurs; instead, the air volume required at the last crosscuts is assumed to enter the mine, travel the various intake air courses intact, sweep the workings and travel out by the return air courses to the point or points of exit from the mine. Actually, there is a leakage of air from intake to return at every stopping, the quantity of this leakage depending upon the tightness of the stopping and the pressure difference across the stopping. Unfortunately, a lack of empirical data and knowledge of the condition of individual stoppings makes an exact analysis of underground stopping leakage impossible. Generally, leakage is most severe through the old stoppings outbye the circuit. These are also subjected to higher pressure differences than the newer inbye stoppings. Therefore the circuit air volume diminishes at a decreasing rate progressing from outbye to inbye the circuit. Consider a simple mine circuit consisting of one intake and one return air course, each of equal and uniform section area and interconnected at regular intervals with crosscuts equipped with stoppings. At any point X inbye the circuit the air volume q on the intake air course equals that on the return air course and decreases with diminishing rate from outbye to inbye the circuit between the limits of Q1 completely outbye and Q completely inbye. Such a volume change is represented by curve a, Fig. I, which is computed with the aid of the equation q = Q1- (Q1- Q)(X + L)N. This equation results from the general algebraic equation of form q = k, — kXN after solving for kl and k with the knowledge that completely outbye X = o and q = QL, also completely inbye X = L and q = Q. Curves a, b and c, Fig. I, show values of q as ordinates versus values of X as abscissas, for values of N equal to 0.50, 1.00 and 2.00, respectively. Distribution of underground stopping 1eakage directly affects the air-volume distribution along the circuit, which in turn determines the circuit pressure gradient. To deliver any required air volume inbye a circuit or portion thereof requires a higher circuit-pressure differential if stoppings leak than is necessary with tight stoppings. Every mining engineer engaged in the analysis of projected mine-ventilation circuits is therefore confronted with the problem of making proper allowance for the effects of leaking stoppings. The purpose of this paper is to present evidence in support of the author's suggestion that circuit pressure increase (per cent)

Jan 1, 1942

By Raymond Mancha

When calculating the pressure-volume characteristics of projected mine-ventilating circuits by orthodox methods, certain basic assumptions are required in order to employ the various available empirical data. It is assumed, for example, that the mine air is an incompressible fluid subjected to isothermal flow, an assumption sufficiently accurate for practical purposes since pressure and temperature differentials are small throughout the average circuit. A more erroneous assumption, however, is that all air is accounted for as it travels throughout the mine. No attempt is usually made to evaluate stopping leakage as it occurs; instead, the air volume required at the last crosscuts is assumed to enter the mine, travel the various intake air courses intact, sweep the workings and travel out by the return air courses to the point or points of exit from the mine. Actually, there is a leakage of air from intake to return at every stopping, the quantity of this leakage depending upon the tightness of the stopping and the pressure difference across the stopping. Unfortunately, a lack of empirical data and knowledge of the condition of individual stoppings makes an exact analysis of underground stopping leakage impossible. Generally, leakage is most severe through the old stoppings outbye the circuit. These are also subjected to higher pressure differences than the newer inbye stoppings. Therefore the circuit air volume diminishes at a decreasing rate progressing from outbye to inbye the circuit. Consider a simple mine circuit consisting of one intake and one return air course, each of equal and uniform section area and interconnected at regular intervals with crosscuts equipped with stoppings. At any point X inbye the circuit the air volume q on the intake air course equals that on the return air course and decreases with diminishing rate from outbye to inbye the circuit between the limits of Q1 completely outbye and Q completely inbye. Such a volume change is represented by curve a, Fig. I, which is computed with the aid of the equation q = Q1- (Q1- Q)(X + L)N. This equation results from the general algebraic equation of form q = k, — kXN after solving for kl and k with the knowledge that completely outbye X = o and q = QL, also completely inbye X = L and q = Q. Curves a, b and c, Fig. I, show values of q as ordinates versus values of X as abscissas, for values of N equal to 0.50, 1.00 and 2.00, respectively. Distribution of underground stopping 1eakage directly affects the air-volume distribution along the circuit, which in turn determines the circuit pressure gradient. To deliver any required air volume inbye a circuit or portion thereof requires a higher circuit-pressure differential if stoppings leak than is necessary with tight stoppings. Every mining engineer engaged in the analysis of projected mine-ventilation circuits is therefore confronted with the problem of making proper allowance for the effects of leaking stoppings. The purpose of this paper is to present evidence in support of the author's suggestion that circuit pressure increase (per cent)

Jan 1, 1942

On Friday, Jan. 17, the thirteenth impression of the Perkin Medal was presented to Dr. Frederick Gardner Cottrell, at a meeting of the New York Section of the Society of Chemical Industry, held at the Chemists' Club, in New York City. The presentation address was delivered by Dr. C. F. Chandler, Senior Past President of the Society. The medal was awarded for the recipient's work in the electrical precipitation of suspended particles in gases. Dr. Cottrell's career, which is so familiar to members of the Institute, was sketched by Dr. Chandler. He told at some length of the experiments leading up to the final solution of electrostatic precipitation, which has done so much to solve the problem of smelter smoke. The principal interest in Dr. Cottrell's speech of acceptance lay in his remarks about helium, the production of which has been so reduced in cost that it has practically supplanted the use of hydrogen for inflation purposes in war balloons. He said: "Two or three weeks after war was declared in 1917, Dr. R. B. Moore attended a meeting of the American Chemical Society in Kansas City. At the general session, Mr. Seibel of the University of Kansas, gave a paper on krypton and xenon in some of the natural gases of Kansas. At the end of his paper, he expressed regret that at such a time, when everyone was thinking of war problems, his paper was of a purely scientific nature and had no practical bearing on the war. Dr. Moore immediately got up and said that he did not agree with Mr. Seibel. The presence of helium in those wells could, and should, have a very practical bearing on the war, as this gas could be extracted in quantity from the natural gas and used for balloons and Zeppelins. He pointed out some of the natural advantages of helium over hydrogen, and quoted Sir William Ramsey's letter advocating the use of helium for airships. The general attitude of those present was one of scepticism, so he finally asked Dr. Cady for his opinion who stated that he believed-the thing could be done, but was very doubtful whether it could ever be made practical on account of the cost." Dr. Moore in a letter wrote: "That same day I talked to Dr. Parsons, who was present at the meeting and was returning to Washington almost immediately. I told him what had happened, and that I believed the matter should be taken up by the Bureau at once and presented to the War Department. He promised to do this as soon as he got to Washington; and I know that he did take the matter up with other Bureau officials." Dr. Cottrell continued in part: News of Ramsey's suggestions also reached this .country through other channels and eventually came to our attention. Professors Satterly and Patterson of the University of Toronto commenced experimentation of the subject Jan. 1, 1916. Col. G. A. Burrell, who headed the research department of the Gas Warfare Service at the American University, tells me that Cady's experiments had also suggested similar possibilities to him.

Jan 3, 1919