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By R. A. Kinzie

In a conventional cement plant, the drying and heating of the raw material takes place in the upper section of a cylindrical kiln where the heat exchange is poor. In a Lepol kiln this part of the process is carried out on a traveling grate where the heat exchange is efficient. The increase in fuel economy justifies the greater complexity of the installation.

Jan 1, 1950

By L. A. Mekler

The recirculating furnace is primarily a heating apparatus of the convection type in which the heat-absorbing surfaces are heated by a mixture of fresh products of combustion and a portion of the combustion gases that have already given off a part of their heat in their previous passage over the heat-absorbing surfaces. In its basic conception, the recirculating furnace can be considered as a closed thermal system in which the heat-absorbing surfaces are constantly swept by a fixed quantity of products of combustion at a given initial temperature with the gases being constantly returned to the furnace and reheated before they again enter the heating zone. Gases equal in amount to those produced by the fuel burned to reheat the circulating gases are constantly ejected from the system to the stack. The recirculating furnace is the most suitable apparatus for heating by convection where low furnace temperatures must be maintained and where only a small portion of the available heat is taken out of the gases, particularly where the maximum allowable furnace temperature is low and where the final temperature of the gases is only a few hundred degrees below their initial temperature. In such cases, recirculation not only provides a method of reducing the rather high combustion temperatures of the fuel when burned without a large amount of excess air, but, by returning to the furnace the heat content of the gases at their final temperature, recovers a large portion of heat that otherwise would have been lost through the stack. The convection-type oil still is particularly suitable for the application of flue-gas recirculation. In an oil still, the heating of the oil is accomplished by applying to the tube surfaces a given quantity of heat of a certain predetermined quality, usually with the initial temperature of the gases not over 1500" F. as measured by thermocouples placed above the tube bank. To obtain the full benefit of the heating surface of the coil, the final temperature of the gases, after passing over the coil, is kept 200" to 300" higher than the temperature of the incoming oil. This final temperature varies from 500" F. for a topping still with only a slight or no

Jan 1, 1928

By Stanly A. Easton, H. M. Childs

Methods evolved at Kellogg have had primarily in view the safety of workmen and complete recovery of the ore; it is believed that these results have been secured at a minimum cost. The ore in the Bunker Hill mine is of good grade, relatively speaking; much of it is quite rich and no mining method would be successful which did not recover all of the ore. It would be possible to work certain of the orebodies by shrinkage methods, or some form of top-slicing, at probably lower costs than by the methods now applied, but the shapes of the orebodies are so irregular that satisfactory extraction of the ore in all its ramifications would be extremely difficult. In past years, much high-grade ore has been lost in blasting, due to the unskilful use of powder in a galena ore, which, by reason of its brittle-ness and great weight, tends to scatter widely. Any method by which large masses of ore and waste are broken together will result in an imperfect recovery of the ore, and whenever large masses are broken, ore and waste will unavoidably fall down together; ore and waste thus become entangled and, although a pile of broken material may seem to have been well picked over, quantities of ore will inevitably be lost. This fact has been shown by persons reworking the filling in stopes that were mined by methods defective in the foregoing particulars. Safety of the men is secured by avoiding large openings, and by working mainly under timber, and at small faces at close range and easily inspected. In former years the principal source of serious accidents was "falls of ground;" it is now over seven years since a serious accident has occurred from this cause. The most prolific source of accidents during recent years has been "timber and tools," workmen being struck by timbers passed into the stope from the level above; this danger has been greatly reduced by compressed-air warning whistles and the fencing in of dangerous areas. The lead-silver ore of the Bunker Hill and Sullivan and affiliated mines is found in bodies of almost every shape, although vertical or nearly

Jan 1, 1925

By J. Versluys

WHEN a well strikes an oil-bearing layer, the oil has a pressure which is generally sufficient to enable it to rise to near the surface (sometimes above the surface). As soon as a well begins to produce, however, the liquid moves through the pores of the reservoir bed and the pressure in the well becomes much lower than the pressure originally prevailing there. At some distance from the well, however, the pressure in the reservoir bed remains unaltered; thus the pressure of the oil has not only to lift the oil, but also to overcome the friction resistance in the pores. The fact that so many oil wells are gushers is a consequence of the energy accumulated in the gas. In gushing the well acts as a gas-lift. A mixture of liquid and gas (the latter partly dissolved in the former) rises vertically from the oil¬bearing layer through a cylindrical casing to the surface. In time conditions alter and the well ceases to gush regularly, then the gushing can be further promoted by inserting a narrower tube in the well and connecting the top of the oil string to the tubing. If the action in time becomes irregular, the gushing can be kept up for a further period by forcing gas between the two tubes. In the oil fields the term "gas-lift" is used actually only where extraneous gas is applied, as in the last of the stages mentioned. The action, however, is just the same whether the gas exclusively originates from the formation, or is partly applied artificially. Thus by gas-lift we simply mean a vertical tube in which the energy of gas under pressure, and of dissolved gas, is utilized for raising a liquid. In gushing oil wells the pressure is frequently very high and the absorption coefficient 0.4 (expressed in vol. ratio) of the coexisting gas is not particularly high, so that in reality it should be assumed that a considerable portion of the gas, at any rate at the bottom of the gas-lift, is dissolved in the oil. For water-producing wells this is not usually of such importance.

Jan 1, 1929

By D. E. Marks, Arnold, C. W, R. J. Robinson, A. E. Hoffmann

The efficiency of operation of a fixed-bed adsorption unit is infEuenced both by the absolute adsorption capacity of the bed and by the rate of adsorption. This paper describer studies of adsorption rate which were conducted in an experimental unit designed such that conditions existing in the treatment of high-pressure natural-gas mixtures could be duplicated. Variables investigated included pressure, temperature, gas composition, adsorbent particle size, depth of packed bed and gas velocity. The adequacy of a simplified mathematical model for predicting the observed phenomena was tested. A correlation is preserited which relates adsorption rate to the process variables stlldied. This correlation is useful in combination with the matheinatical model. INTRODUCTION Of the techniques available for contacting adsorbent particles with fluid streams to be treated, fixed-bed adsorption columns offer definite advantages in simplicity and ease of operation. As a result, they are often used in preference to others for such petroleum industry applications as dehydration and purification of natural gas and hydrocarbon recovery. Fixed-bed adsorption units usually consist of two or more towers filled with a desired adsorbent and operated in a cyclic manner. While one is being used to process the main flow stream, the others are undergoing regeneration to remove the adsorbed phase. When the tower on stream becomes saturated with the preferentially adsorbed material, the roles of the towers are switched, and the freshly regenerated tower is placed on stream. Cacle duration is determined by the bed capacity under the process conditions and by the flow rate through the bed. The sharpness of separation which can be effected is a function of both the absolute capacity of the bed and the rate of adsorption in the bed. The effect of rate for a particular set of conditions is evidenced by the sharpness or diffuse-ness of the adsorption front as it advances through the bed. Since data needed for design of adsorption units to treat high-pressure natural-gas systems were not available, an experimental program was designed to investigate the effects of different variables upon adsorption rate in fixed beds. In the present paper, effects of gas composition, column length, temperature, pressure, adsorbent particle size and flow rate (actual linear flow rate of the gas) are shown, and utility of a simplified mathematical model for describing the process is discussed. As gas enters the top of a cool, clean bed of adsorbent, preferentially adsorbed materials are stripped from the main flow stream by the uppermost particle layers. As these layers become saturated with a particular component, new supplies of this component are carried further down the column until fresh adsorbent is encountered. An adsorption wave thus moves through the column as material is supplied to saturate succeeding elements of the bed. Adsorption from a Multicomponent gas stream occurs as a succession of such moving waves corresponding to the different components in the gas. The leading edge of an adsorption wave for a component of a natural-gas stream moving through a bed of a common commercial adsorbent such as silica gel would be sharp but for the influence of certain broadening fac tors. These factors include a nonuniform velocity profile in the bed, longitudinal dispersion or mixing in the main gas stream, and the time required for a molecule to migrate from the main gas stream and be adsorbed at a site within the body of an adsorbent particle. If packing is uniform and the ratio of column to particle diameter is greater than approximately 15:1, the first factor is relatively unimportant' Longitudinal mixing is of importance only for the case of moderately high mass transfer with extremely slow flow rates.' The sharpness of an adsorption front, therefore, is, primarily a function of the rate of adsorption or the time required to saturate a particle of zdsorbent. Two methods for defining adsorption rate are used in this work. The first is a normalized or relative rate which describes the rate of saturation of a differential element of the packed bed. This can be measured by observing the time required for the concentration of the preferentially adsorbed material in the effluent gas from the bed to rise from zero to a value equal to that in the inlet gas stream. The second definition describes the absolute rate of mass transfer from the gaseous to the adsorbed phase. This definition is used in a mathematical description of the adsorption process. If the concentration of a component in the gas strcam leaving an adsorption column is measured and plotted as a function of time, a curve such as that shown in Fig. I results. It is seen that for a period of time the effluent gas is devoid of the component under consideration. As the bed approaches saturation, a small percentage of this material will appear in the effluent gas. The concentration will then rise with time, or increasing cumulative gas flow, until it is equal to that in the inlet gas stream. If adsorp-

By Allan J. Clark

THE September issue of MINING AND METALLURGY might almost have served as a review of the advances in the metallurgy of gold during the current year. In addition to a scholarly article by F. W. Bradley, there were able descriptions by W. W. Bradley and W. F. Boericke of the methods followed in California and in Ontario. Considered together, the two last-named articles indicate the many possible combinations of flotation with one or more of the older processes. This, the prominence of flotation as an aid in the beneficiation of gold ores, is probably the most striking of the year's developments. It would perhaps be nearer the fact to say that the year has not been marked by the gold metallurgists' discovery of the flotation process, but rather by the flotation merchandizers' discovery of gold metallurgy. I say this because native gold was floated successfully more than fifteen years ago, and a reference to these results was published six years ago.

Jan 1, 1933

By Howard Eavenson

THE eleven mines of the United States Coal and Coke Co. in the Pocahontas coal field are situated in McDowell County, W. Va., which is a mountainous region. The valleys rarely exceed 200 ft. (60 m.) is width, the hill sides slope from 25° to 33°, and the tops, from 600, to 900 ft. above the valleys, carry practically no level land. At least 80 per cent. of the area is covered with timber. All the workings are drift mines; and while the average grades of the seams worked are ample for drainage, local dips and swamps necessitate, the use of a large number of local pumps or of numerous drainage ditches to keep the mine free from water during heavy rains. The water handled in any mine is not great, compared with the coal output, and usually causes little trouble except during extremely wet periods. In hew workings, the water encountered, except in rare cases, is trifling; it is only after rib drawing has started and the surface has been broken that water troubles begin. The strata overlying the coal are sandstone, shales, and slates; as the soil is thin and sandy, it does not tend to puddle and stop the cracks, which remain open for a long time. As mining progresses and the robbed area increases, the broken area becomes larger, increasing the openings into the mined territory. The cover over the robbed portions of the mines varies from 30 to 750 ft. (9 to 228 m.). When the ground is saturated, considerable water finds its way into the mines the day after a heavy rainfall and for three or four days it tends to fill the low places, wash the tracks, etc. and. unless from three to four times as many pumps are installed as are normally needed, considerable', output is lost. In most cases the water will .flow from the mines, so the obvious solution is to provide drainage .ditches of ample size to carry the maximum flow expected, so that the inflow will have no opportunity to flood the working places.

Jan 1, 1921

By Fred H. Poettmann, J. K. Welchon, A. F. Bertuzzi

A series of gas-lift tests was made which verified conclusions reached from previous studies and which showed that gas-lift performance could be calculated if reasonably accurate data on the producing characteristics of the well are available.' The test installation is described. Its essential features consisted of a rate-of-flow controller to regulate the input gas and an auxiliary string of small diameter tubing to conduct the input gas to the point of injection in the eductor tubing. In general, it was shown that for continuous gas-lift operation, injection should be as deep as possible consistent with available gas pressure and rate of production desired. Calculated curves are presented which completely characterize the gas-lift performance of the well tested. INTRODUCTION A correlation based on field data from a large number of flowing and gas-lift wells, covering a wide range of oper- ating conditions, was developed which permitted the logical design and prediction of the performance of continuous gaslift installations.' The procedure made it possible to calculate the depth, pressure, rate at which to inject gas, horsepower requirements, and the effect of production rate and tubing size on these quantities. As a result of these calculations certain significant facts were brought to light or re-emphasized. The presence of efficient ranges of operation and an optimum injection depth for a given rate of production were shown to exist as well as the fact that the lower the tubing pressure the less the horsepower requirements to lift the reservoir fluids. The fact that the gas lift performance can be predicted by calculation does not mean that any given installation will perform over the entire range predicted. Many of the conventional continuous gas-lift installations will not permit operation over the entire range predicted by calculation (particularly in the efficient range of operation) because of the performance characteristics of the installation. Consequently, field tests were initiated for the purpose of investigating and arriving at a simple continuous gas-lift installation which would operate under all conditions predicted by calculation. The first series of tests involved the use of mechanically operated valves. These valves were slip-joint type valves which are opened and closed by lowering or raising the tubing at the surface by means of a hydraulic jack. These valves were chosen because it was hoped that they could be opened or closed at will and did not depend on the gas pressure in the annulus or the pressure of the fluid in the tubing for their operation. Unfortunately, the valves did not operate satisfactorily and it was imossible to open or close each valve at will. Some limited data were obtained, however, which verified the predicted benefits resulting from the use of lower tubing pressures. In the next series of tests the valves were removed and the gas was injected from the annulus through a 14/64-in. orifice equipped with a reverse check valve. A packer was set below the valve to seal off the space between the 2-in. tubing and 5-in. casing. The injection point was located at 3,810 ft. This arrangement provided satisfactory mechanical operation. However, it was impossible to obtain low rates of liquid production at low tubing pressures as predicted by calculation. A minimum gas injection rate existed below which steady flow was impossible. Liquid pro-

Jan 1, 1953

By Orville R. Lyons

CENTRIFUGAL force has been utilized for the dewatering of fine coal for over 50 years by means of machines commonly called centrifugal dryers. In any centrifuge the coal and water are subjected to a spinning action which usually inctedtocreases in intensity as the coal progresses through the machine. This spinning action, or the centrifugal force that it induces, tears water away from coal particles and produces a dewatered coal. The effectiveness of the dewatering action for any particular machine is governed by the size-consist of the coal and the centrifugal force imparted to the water on the coal. Four makes of centrifuge are currently being used to dewater fine bituminous coal in the United States. These include the vertical, basket-screen C.M.I. (Centrifugal and Mechanical Industries) Carpenter, Reineveld, and the horizontal solid-bowl Bird. The C.M.I. and Reineveld centrifuges are similar in construction and are variations of the Elmore centrifuge, varying principally in the slope of the basket, operating speed, size of perforations in the basket, and diameter of the basket. The same operating description suffices for both. These centrifuges are made up of two rotating elements, an outside conical screen frame and an inside solid cone which carries spiral pusher blades. The screen frame is supported by a hollow shaft, and the solid cone by a spindle shaft passing through the hollow shaft. Both the screen frame and the solid cone rotate at high speeds; the solid cone, with its scraping flights, rotating somewhat more slowly than the screen. The wet coal enters the machine at the top, falls on the cone, and is thrown against the screen by centrifugal force. The coal slides down the screen until it meets the upper end of the flights, where it is moved slowly along the flights until discharged at the bottom. The elfluent passes through the sieve basket or screen and discharges into a launder. The Carpenter centrifuge consists of a conical rotating element with a vertical axis, built up of three rows of perforated stainless-steel plates or stainless-steel wedge-wire or round-wire sections cut and rolled to conform to the surface of the cone. Wet material is delivered to a conical hopper inside the centrifuge casing and is distributed in a thin, even layer over the inner surface of the cone by a disk mounted on the rotor shaft. Impacting of the coal against the screen and the centrifugal force exerted as the material moves downwards creates a dewatering effect. The moisture is forced through the coal bed and then through the perforated plates of the basket, collecting in a circular trough built into the centrifuge. The effluent is conducted from the trough through the centrifuge casing by outlet pipes. The dewatered coal is discharged over the inside bottom edge of the rotor. The Bird centrifuge consists of a tank or truncated conical shell, which is revolved at the desired speed by means of a drive sheave. A screw conveyor rotates inside the cone or bowl at a slightly lower speed in the same direction of rotation. The feed entrance, in the center of the large end of the truncated cone, is high enough to allow formation of a pool of slurry. Adjustable effluent-discharge ports are located in the large end of the bowl so the level of the liquid can be regulated. The solids are moved forward by the screw conveyor as fast as deposited and carried above the level of the pool for an interval before leaving the bowl. In the past these four makes of centrifuge could be compared only by indirect means. The author published a rough comparison method in 1949' but

Jan 1, 1952

By K. D. Sheffler, R. W. Hertzberg, R. W. Kraft

The Ni-Ni,Ti lamellar eutectic alloy responds to unidirectional solidification by alig-nment of the platelets of the two phases roilghly perpendicular to the moving solid-liquid interface. X-ray diffraction and zetallographic experimenls show that the following cryslallographic relationships exist in the directionally solidified ingots: lamellar interface w [0001] Ni3Ti 11 (111 }Ni local growth direction (1120) Ni3Ti n (110) Ni These relationships are rationalized on the basis of the crystallography of the constituent phases. In recent years researchers in several laboratories have shown that an aligned or "controlled" micro-structure can be produced in many binary eutectic alloys by forcing the alloy to solidify unidirectionally in an elongated crucible.'-3 It also has been found that a unique crystallographic relationship often develops between the two phases in such a structure.4, 5 Since a parallel arrangement of phase particles in an alloy should, and, in many cases, does, produce unique and highly anisotropic properties, the alloys so produced are of potential interest in a variety of applications. Studies of these alloys are also of interest at the academic level because they help to improve our basic understanding of the way in which metals solidify. This paper reports the metallographic and crystallographic structure of another of these alloys—that between nickel and Ni3Ti. In a separate paper, hot tensile and stress rupture properties as well as deformation and fracture behavior of the same alloy are discussed.19 EXPERIMENTAL DETAILS AND RESULTS General Characteristics of the Eutectic Reaction. Several small exploratory Ni-Ti melts were made to resolve a conflict in Hansen6 over the eutectic composition. The alloy containing 16.4 at. pct Ti, Fig. 1, appeared metallographically to be closest to the eutectic composition, although small dendritic patches of proeutectic nickel are present. The eutectic morphology is lamellar and appears quite sharp and angular, indicating a strong influence of crystallographic factors upon the microstructure. Some of the eutectic nickel platelets are continuous with the primary phase, whereas in hyper eutectic melts a "halo" of nickel generally surrounded the proeutectic inter-metallic platelets, indicating that nickel is the nucleating phase in this eutectic reaction. Extensive precipitation of Ni3Ti has occurred in the primary nickel phase in a Widmanstatten pattern which makes the dendrites difficult to differentiate from the eutectic mixture. This phenomenon may account for discrepancies among the eutectic compositions reported in the literature. The sloping solvus line bounding the nickel primary solid solution explains the presence of the intermetallic precipitate. Poole and Hume-Rothery7 report that the solubility of titanium in nickel drops from 13.8 at. pct at the eutectic temperature (1304°C) to approximately 9 at. pct at 700°C (Ni3Ti, on the other hand, melts con-gruently and exhibits essentially no solubility for nickel or titanium). Lever rule calculations based upon Poole and Hume-Rothery's phase diagram indicate an increase from 21 vol pct Ni3Ti in the eutectic mixture at the eutectic temperature to 61 pct at room temperature. Selective point count measurements of the volume fraction of eutectic Ni3Ti platelets in the eutectic microstructure yielded a value of 46 i 2 vol pct, which required extensive precipitation of Ni3Ti from the nickel solid solution onto the eutectic intermetallic platelets during cooling from the solidification temperature. Influence of Directional Solidification. Nickel bar and titanium sponge of 99.9 pct purity were induction-melted in zirconia crucibles under argon in the proportion of 13.8 wt pct (16.4 at. pct) Ti to produce 14-lb master heats which were cast into investment molds yielding 5/8-in.-diam pins approximately 10 in. long. These pins were remelted under argon in horizontal zirconia boats and unidirectionally solidified to produce 1/2-in.-sq ingots approximately 44 in. long. Details of the experimental techniques and apparatus are described elsewhere.'

Jan 1, 1970

By Brown, R. W., G. H. Neill, R. G. Loper

All facets of ball sealer behavior must be known and understood to design for their optimum use in well treatments. The down-hole factors including the inertial forces, drag forces and holding forces have been discussed. Equations have been developed and presented for these forces. The use of these equations in this process will contribute to the most efficient use of an economical and proven ball sealer process. The urocedure presented has been used to successfully analyze field results indicating problems involving the use of ball sealers. These problems have been isolated and corrected in future treatments by using the procedure presented to revise the perforating program, the injection rate, the design of ball sealer stages and/or the type and size of ball sealers. The procedure presented also indicates that dislodgement of ball sealers occurs primarily in ultra slim-hole completions. INTRODUCTION The perforation sealing process has been proven highly successful and economical since it was introduced to the oil and gas industry in early 1956.' Since it has proved so successful in multistage fracture treatments at a much lower cost than conventional packers, the ball sealer process has become a byword in well completions.' Not only has it changed the design of fracture treatments, it has changed the concepts of selectively perforating and has greatly aided the success of the single-point and limited-entry techniques." Yet with all the advantages, there are still facets of ball sealer behavior that are not widely recognized. These must be known and understood to better design and execute the optimum ball sealer treatment. The down-hole behavior of ball sealers is influenced by a number of factors including an "inertial force", a "drag force" and a "holding force". These are considered in two phases—the initial seating of the ball and the forces tending to unseat the ball after contact has been made (see Fig. 1). This paper presents a theoretical study of these factors and formulas for predicting their influence. The authors acknowledge the fact that the flow equations presented herein describe the behavior of Newtonian fluids and do not cover the so-called power-law fluids. The assumptions made, however, are reasonable and well within practical limitations of the conditions encountered during field application. For example, it is assumed that the perforations are round with a discharge coefficient of 0.82. While this may not be exactly true, it has been verified within practical field limitations.' The discharge coefficient would not vary significantly even if the perforation size or configuration varied pronouncedly. The approach of assuming uniform perforation size is not as strained as it might first appear. Normally, a perforator contains charges equally spaced with the same number in each vertical (or horizontal) plane. If the perforating gun is not centralized, the holes in one plane will be smaller than average, while the holes in opposing planes will tend to be larger than average. The net result is that the two effects tend to balance. FACTORS AFFECTING CONTACT WITH PERFORATlON Ball sealers are injected into well treating fluids for the purpose of contacting and sealing those perforations which are accepting the fluid flow. The efficiency of the sealers is primarily influenced by: (1) the velocity of the balls down the pipe, and (2) the fluid velocity through the perforations. To divert the sealer to the perforation, the inertial force of the ball must be overcome by the drag force created by the fluid velocity through the perforation. BALL VELOCITY The final, or stable, velocity of the ball is the sum of the actual fall velocity and the fluid velocity. This will be

By R. S. Rickard, M. C. Fuerstenau

Anodic polarization curves for iron dissolution and cathodic polarization curves for copper deposition and ferric and hydrogen ion reduction were studied. These results were used to predict the relative rates of the reactions: The rates of the first two reactions were found to be first order with respect to the reacting ion, and the rale-controlling step was diffusion of the ions to the callzode surface. The rate of hydrogen ion reduction in an oxygen-free solution is dependent on the reciprocal of the square root of the hydrogen ion activity. In solutions containing oxygen, the rate of hydrogen reduction becomes dependent on the square root of the partial pressure of oxygen. At equal concentrations copper precipitation occurs about twice as fast as ferric ion reduction. In an air-saturated solution containing 0.01 M cupric ion at pH 2, the rate of hydrogen ion reduction zs about 10 times slower than that of copper precipitation. At pH values below 3, the rate of hydrogen reduction in oxygen-free solutions is very slozu. COPPER cementation has been utilized industrially for many years as a means of recovering copper from aqueous solutions. There are many different types of cementation plants, but all simply provide a means of contacting iron with relatively dilute copper solutions. In the past little or no attention has been paid to rates or mechanisms of the reactions involved. As a result, this and other studies'-3 have been undertaken recently to establish the important parameters in this system. The reactions involved in copper cementation by iron are: Since transfer of charge is involved in every case, these reactions should be considered from an electrochemical point of view. In any electrochemical system involving electrolytic conduction, passage of electrons necessitates ion transfer to the respective electrodes according to Faraday's law. The current in the system will be proportional to the rate of ion transfer, and, therefore, when the current in the galvanic cell is known, the rates of the reactions involved can be determined. Galvanostatic electrode measurements5" are applicable for these electrochemical studies, in that with this technique polarization curves (current density vs electrode potential) are established by impressing a constant current through an electrode and measuring the electrode potential. Once the current density, electrode potential, and reacting area are known for each reaction in this cementation system, the rates of the iron-consuming reactions can be determined. EXPERIMENTAL MATERIALS Several metal electrodes were used in this investigation; they are shown in Table I. Solutions were made with reagent-grade salts and conductivity water. Air, nitrogen, and oxygen were employed. All of these gases were saturated with water and degreased prior to use. EXPERIMENTAL PROCEDURE The experimental work was divided into two general areas. The first area involved galvanostatic measurements, which were used to determine the rates of reactions as a function of various parameters. The second area comprised actual cementation experiments to check the rates predicted from the electrode measurements. Galvanostatic Electrode Experiments. Polarization curves were obtained using a standard galvanostatic technique5'= with the apparatus shown in Fig. 1. The various electrode polarization experiments were run the same way except for changes in solution composition, polarity, and type of measured electrode. The general procedure involved the following steps: 1) a solution was made 0.5 M in sulfate ion and to the desired cation concentrations; this solution was placed in a constant-temperature bath and saturated with the desired gas; 2) the electrodes were cleaned and placed in the electrode holder; 3) the previously cleaned apparatus was assembled and immersed in a constant-temperature water bath (temperature controlled to ± 0.05"C); 4) the electrode was then positioned; the solutions were introduced, and the polarization meas-

Jan 1, 1969

By A. W. Schlechten, C. M. Hsiao

The apparent vapor pressures of a number of metal sulphides were determined by measuring their rate of weight loss when they were heated under vacuum. The calculated pressures are due in some instances to the volatility of the sulphide and in others to the decomposition pressure. FIE direct reduction of metallic sulphides, as in the precipitation method for lead or antimony, is not a widely used process today. However, it might be employed for the production of a number of metals if the operation were carried out under reduced pressures. Two general procedures are possible; the reducing agent could form a stable nonvolatile sulphide from which the valuable metal could be removed by volatilization or liquation. The other approach would be to choose a reducing agent which would form a sulphide that could be volatilized away from the valuable metal. A vacuum operation would serve to protect the sulphides and metals from oxidation and would increase the evaporation rates. A related process is the separation of metallic sulphides by selective volatilization either at atmospheric pressure or under vacuum. This procedure is already used commercially for the elimination of lead and cadmium as PbS and CdS from zinc calcines; it might well have other applications. To predict such processes it is necessary to know the relative volatilities of the metallic sulphides and to know whether they volatilize as sulphides or whether the predominant weight loss is due to decomposition. Such data would also be of value in studying problems such as the use of sulphides for refractories, as described by Brewer,] or in the evaluation of the theory of ore deposition from the vapor state, as proposed by Brown.' Qualitative statements concerning the volatility of metallic sulphides can be found frequently in the literature and are well summarized by Mellor, but accurate determinations of rates of evaporation or vapor pressures are very scarce. Kelly" lists vapor pressure data only for PbS; his reference being to the work of Schenck and Albers,4 who measured the vapor pressure of PbS by a dynamic method between 850" and 995°C. More recently Veselovskii5 has determined the vapor pressures of the sulphides of antimony, lead, cadmium, and zinc at temperatures ranging from 360" to 1000°C using Knudsen's method. A second Russian worker, Pogorelyi,6 determined the vapor pressures of CdS and ZnS at temperatures from 900" to 1250°C. The results obtained by these workers are shown in subsequent graphs in comparison with the data of this paper. Brewer and coworkers1 have studied some of the metallic sulphides to determine their suitability as refractories in vacuum. Veselovskii's determinations on PbS checked well with those of Schenck and Albers, and his determinations of CdS check almost exactly with Pogo-relyi. For the vapor pressure of ZnS the data of Veselovskii and that of Pogorelyi are in fair agreement, as shown in Fig. 1, but it should be pointed out that Lumsden7 has discovered that the equation given by Veselovskii for the vapor pressure of ZnS: - 14,200 log P9.495 is in error by a factor of ten when compared to the experimental data. In contrast with the scarcity of data on vapor pressures, the decomposition pressures of many metal sulphides have been determined and are summarized by Smithells.8 Theoretical Considerations The experiments described in this paper consisted of determinations of the rate of weight loss of certain metallic sulphides under very low pressures at different temperatures. These data plus observation and analyses of the products reveal the mechanism of the weight loss and also provide a basis for the calculation of the vapor pressure or apparent vapor pressures of the sulphides.

Jan 1, 1953

By J. L. Carne

ORE control is a matter of planning and supervision based on a foreknowledge of the content and distribution of ore. The Inspiration orebody is predominately a copper-sulphide blanket, overlain by an oxidized zone of copper silicates, copper carbonates, and barren capping. Most of the ore is treated by leaching, and the optimum requirement of this method is the goal toward which the mining schedule is aimed. Both open-pit operations and underground mining are used. The problems of ore control at Inspiration are: I—maintenance of metallurgical balance, 2—conservation of oxide ore, and 3—mining out of the orebody at an average grade which parallels that of the reserve figure. Prior to the fall of 1926 all of the Inspiration division ores were treated by flotation, but the mixed ores of the Live Oak—Keystone section carried too large a proportion of copper in the form of chrysocolla for satisfactory flotation. For this type of mixed ore the leaching process now in use had been worked out, and the plant was put into operation in the fall of 1926. The process is known as the ferric sulphate leach, and the leaching solvents contain both ferric sulphate and free sulphuric acid. Sulphide copper in the form of chalcocite is leached by the ferric sulphate, and the oxidized copper minerals are dissolved by the sulphuric acid. From the beginning of operations to the present time the plant has treated some 67 thousand tons of ore having an average copper content of 1.142 pct, of which 0.605 pct has been in the form of oxidized copper minerals. The plant is currently being operated at the rate of 11,500 tons per day, 7 days per week. However, this tonnage is produced in 6 days of mine operation. The average current grade is approximately 1 pct copper, of which about 50 pct is present as oxidized or acid-soluble copper. Close control of metallurgical procedures is a vital factor in operation of the plant and is in turn dependent upon careful regulation of feed. An excess of sulphide copper requires the presence of more ferric sulphate in the leaching solvents. On the other hand, an excess of ferric sulphate in the presence of a low sulphide feed seriously reduces the efficiency of electrolytic precipitation in tank house operations. This condition must be avoided at all costs. An additional factor must be considered. It is known that the sulphide content is predominant in the ore which remains in the reserve. To maintain the necessary sulphide-oxide proportion, so that the entire reserve will mine out at both a grade and a sulphide-oxide ratio that will be permissible throughout the remaining life of the property, it is obvious that the oxide reserve must be currently preserved to as great an extent as possible. Fortunately much of the oxide reserve is found in the open-pit operation. This happy circumstance adds flexibility to the procedure and makes it possible to maintain the close regulation necessary to insure best overall results, both mining and metallurgical. The two ore streams from the pit and underground operations join at the coarse-crushing plant, and proper control must be exercised before the ore reaches this point. The open pit is currently providing about 55 pct of the ore and underground operations the balance of production. Both operations are under a single superintendent so that the closely interrelated problems may be more readily controlled. The least flexible of the two producing units is the underground mine, where the mining method is block caving. Although a block of ground may be undercut and developed, characteristics of the material determine its caving action. The ore in the chutes must be drawn. The area currently mined from underground is in a zone predominately sulphide and is generally

Jan 1, 1953

By J. L. Carne

ORE control is a matter of planning and supervision based on a foreknowledge of the content and distribution of ore. The Inspiration orebody is predominately a copper-sulphide blanket, overlain by an oxidized zone of copper silicates, copper carbonates, and barren capping. Most of the ore is treated by leaching, and the optimum requirement of this method is the goal toward which the mining schedule is aimed. Both open-pit operations and underground mining are used. The problems of ore control at Inspiration are: I—maintenance of metallurgical balance, 2—conservation of oxide ore, and 3—mining out of the orebody at an average grade which parallels that of the reserve figure. Prior to the fall of 1926 all of the Inspiration division ores were treated by flotation, but the mixed ores of the Live Oak—Keystone section carried too large a proportion of copper in the form of chrysocolla for satisfactory flotation. For this type of mixed ore the leaching process now in use had been worked out, and the plant was put into operation in the fall of 1926. The process is known as the ferric sulphate leach, and the leaching solvents contain both ferric sulphate and free sulphuric acid. Sulphide copper in the form of chalcocite is leached by the ferric sulphate, and the oxidized copper minerals are dissolved by the sulphuric acid. From the beginning of operations to the present time the plant has treated some 67 thousand tons of ore having an average copper content of 1.142 pct, of which 0.605 pct has been in the form of oxidized copper minerals. The plant is currently being operated at the rate of 11,500 tons per day, 7 days per week. However, this tonnage is produced in 6 days of mine operation. The average current grade is approximately 1 pct copper, of which about 50 pct is present as oxidized or acid-soluble copper. Close control of metallurgical procedures is a vital factor in operation of the plant and is in turn dependent upon careful regulation of feed. An excess of sulphide copper requires the presence of more ferric sulphate in the leaching solvents. On the other hand, an excess of ferric sulphate in the presence of a low sulphide feed seriously reduces the efficiency of electrolytic precipitation in tank house operations. This condition must be avoided at all costs. An additional factor must be considered. It is known that the sulphide content is predominant in the ore which remains in the reserve. To maintain the necessary sulphide-oxide proportion, so that the entire reserve will mine out at both a grade and a sulphide-oxide ratio that will be permissible throughout the remaining life of the property, it is obvious that the oxide reserve must be currently preserved to as great an extent as possible. Fortunately much of the oxide reserve is found in the open-pit operation. This happy circumstance adds flexibility to the procedure and makes it possible to maintain the close regulation necessary to insure best overall results, both mining and metallurgical. The two ore streams from the pit and underground operations join at the coarse-crushing plant, and proper control must be exercised before the ore reaches this point. The open pit is currently providing about 55 pct of the ore and underground operations the balance of production. Both operations are under a single superintendent so that the closely interrelated problems may be more readily controlled. The least flexible of the two producing units is the underground mine, where the mining method is block caving. Although a block of ground may be undercut and developed, characteristics of the material determine its caving action. The ore in the chutes must be drawn. The area currently mined from underground is in a zone predominately sulphide and is generally

Jan 1, 1953

By C. BARING HORIVOOD

(Chattanooga Meeting, October, 1908.) THIS paper is intended, in the light of recent investigations, to call attention to some of the essential features of good practice in sampling and mine-valuation. Mine-sampling may be divided broadly into two classes : (1) development-sampling and computation of ore-reserves; and, (2) stope-sampling and estimation of stope-values. DEVELOPMENT-SAMPLING. It is with development-sampling only that this paper deals-the term development being taken to mean the opening-up of a mine by cross-cuts, drives, connections, etc., preparatory to actual extraction of the ore for milling. In development-sampling the work of the sampler and the consequent estimation of the probable value of the mine or portions of the mine, as the case may be, cannot be verified until a later period, when the ground thus sampled has been exploited; hence, extreme care and accuracy are essential. Moreover, the development-sampler has to foresee and allow for contingencies which may arise when stoping is started. It follows that the sampling, in development-work, demands familiarity with the conditions likely to be encountered later in actual mining. In stopes, on the other hand, sampling is clone under actual working-conditions, and the sampler's valuations are immediately checked by the gold-output, or in some cases by a second sampling of the broken ore. A stope-sampler therefore need not be an experienced miner, though he should possess sound technical training and work accurately. Should he be somewhat inexperienced, be soon has his errors discovered for him, and learns to avoid them. Taking these facts into account, it is obvious that a mining company employing two or more samplers should

Jan 1, 1909

By AIME AIME

AT a recent meeting of the Utah Section. A.I.M.E., P. D. Nielson, general plant superintendent of the new Geneva steel plant at Provo, Utah, spoke on "General Operations of the Geneva Plant." Mr. Nielson explained that the plant was located at Provo, Utah, on account of railroad facilities, the proximity of coal, iron, and limestone, and the nearness of water for operating the plant. Water probably was the most important item in making the decision.

Jan 1, 1944

By Theron Wasson

Discoveries in Michigan, which at the beginning of the year 1934 indicated possible new areas, did not develop into fields of market-breaking proportions. Hart, Oceana County, developed small production from the Traverse lime. Edmore, Montcalm County, proved a small area for oil from the Traverse. West Branch, Ogemaw County, previously reported as Ogemaw, had quite a drilling campaign but was somewhat disappointing because wells were small and erratic. During the year, 30 wells were completed at West Branch, and at the end of the year 15 others in and around this field were under way. At West Branch the Traverse lime, found at 1800 ft., has thin porous streaks near the top, which make small wells; the Dundee lime at 2800 ft. produces from a porous zone of varying thickness. Over the entire field the Dundee wells average 10 to 20 bbl. per day. Acid treatment has materially increased production from some Dundee wells in this field. About 3000 acres have been proved productive from the Traverse and Dundee. At the close of the year the well in Beaverton township, Gladwin County, is probably the most discussed of any in Michigan and may be opening a new field of some proportions. It had an initial production of 70 bbl. daily from the Dundee lime at 3830, which has decreased to 25 bbl. per day. Deeper drilling into the lime may increase production. During 1935 the area around this test will probably be the most active in the state. Structural conditions in Gladwin County, as in other parts of Central Michigan, are obscured by a thick mantle of glacial drift and it will take considerable drilling to outline subsurface conditions accurately. The most important new shallow gas area of the year is around Edmore in Mecosta and Montcalm counties. Eight wells have been drilled to the Michigan stray and the upper Marshall sand, having a combined initial production of 30 million cubic feet. These gas wells are scattered over so wide an area that an estimate of proven territory cannot be made until more drilling is done. At the present time no gas is being sold from Edmore.

Jan 1, 1935

By H. E. Mussey

WHEN the American Institute of Mining Engineers visited the Birmingham district in May, 1888, the four Ensley furnaces (Fig. 1) then FIG. 1.-BLAST-FURNACE DEVELOPMENT IN ALABAMA. completed were referred to as monumental.1 Their dimensions are given in Table 1. Mr. Gordon stated, with considerable satisfaction, that on the preceding day one furnace had produced 165 tons. The blast was

Jan 10, 1924

By Arthur A. Center

AT the beginning of 1944 it was expected that the production of metallic zinc in the United States from domestic and foreign concentrates would exceed the 1943 figure though domestic production of concentrates was expected to drop below the 1943 rate, mostly in the Tri-State area. Imports of concentrates were estimated to more than take care of such deficiency. North American concentrate production, available for import to the United States or elsewhere during the year, was estimated at over 200,000 tons recoverable zinc content; South American at about 65,000 tons; Australian at about 50,000 tons. Production of both concentrates and metal in the United States, however, did drop, attributable in considerable part to man-power shortage in the industry. However production exceeded consumption, which fell off, and stocks of metal increased. These stocks have increased since the middle of 1942 but production and consumption are coming nearer to a balance. The A, B, and C premiums, which have played an important part in keeping up domestic production of concentrates, extend only to the middle of 1945 by present legislation. Congress is being asked for their continuance during the war and

Jan 1, 1945