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By Frank J. Tolonen, Nicholas H. Manderfield, Paul Jasberg

EARLY in the study of the low grade iron formations of Michigan, wide variations in their structure and texture became evident. Because of these variations no simple method of concentration is possible, and those portions of the formations that can be exploited profitably under a given stage of metallurgical progress and existing economic conditions must be searched for carefully. Both structure and texture of the formations have an important bearing on their amenability to bene-ficiation. By structure is meant the banding planes of easy fracture, the porosity, and the degree of leaching. Texture, which includes grain sizes, shapes, and degree of interlocking, determines the amount of grinding necessary to liberate the mineral grains. Structure determines the liberation of portions richer in one mineral. Structure and texture may be termed the gross features of the formation. Earliest surveys of the formations were conducted to outline the parts that could be concentrated at % in. (3 mesh). Operators agreed on this size because anything finer required sintering or other agglomeration processes. At 3 mesh only the structural features, as defined above, would be liberated. With suitable ores concentration was easy, as any of the gravity methods could be used. Sink-and-float gave good results, but with this method only limited portions of the iron formations yielded a desilrable product.1,2 n general, the iroa content could be raised to shipping grade, but the silica content remained too high—sometimes 20 pct—instead of 9 or 10 pct. Grinding did not greatly increase the liberation ofsilica until 2010 mesh or finer was reached. To find the reason fpr this T. M. Broderick began micros~opic examin'ations of the formations in 1933.3 Procedure: The purpose of this discussion is to correlate the appearance of typical portions of the iron formations with their amenability to concentration. As liberation is essential before separation becomes possible, much of the work is based on measuring the liberation achieved by crushing and grinding. Both heavy media and magnetic tube tests were used in this work. On the basis of general structural features such as banding, degree of alteration, and leaching, a specimen was selected from each part of the iron formations studied. The specimen represents the gross features of the formation and not necessarily the particular formation quantitatively. One half of the specimen was polished for macro-study, and the remainder was used for analyses to obtain the grades represented by variations appearing in the macrosection. Specimens for microscopic examination were also prepared for each variation in gross appearance. For correlation of microstructure with liberation, the previous work by sink-and-float methods was supplemented by magnetizing roasting followed by magnetic separation. The results show the liberation of silica to be very important at the sizes used, because any silica particle with iron minerals occluded would be held in the concentrate. A magnetizing roast changes the crushing and grinding character of the sample. For partly leached soft samples the change is slight, but with harder types and those in which there is a considerable change in crystal structure of the minerals, the increase in liberation of the iron minerals may be 30 pct more than that of unroasted samples at the same crusheH size. Gogebic Range Iron Formations: The Gogebic Iron Range is located in the extreme western and southern parts of Michigan's Upper Peninsula, extending westward into Wisconsin. Samples for this study were obtained from the Ironwood formation, comprising three chief members—the Plymouth, Norrie, and Ahvil—with several intermediate slates and thick overlying slates. (The slates are not considered in this discussion because grinding finer than present practice is usually required to liberate the iron minerals.) The predominant iron mineral is hematite, about half of which is hydrated to goe-thite. Limonite occurs in some of the partly leached areas, and siderite in parts of the formation.

Jan 1, 1958

By G. R. Hodgson, M. P. Lebourg

The introduction in the field of a new type well completion called for the setting of tubing open-ended in the well before perforating the casing. This paper describes a new perforating tool of the shaped charge expendable type, small enough to pass through tubing and powerful enough to perforate the casing. The perforations are produced at an angle and are large enough for optimum production. Performance of this tool in targets and under various well conditions is described. Very often no lifting equipment will be available at the well. In some cases, pressure will exist in the well at the time of the perforating operation (workover wells or multiple trip perforation) ; therefore, new equipment and new technique for the perforating operation were devised. This paper describes this equipment and technique, particularly casing collar recording, sheave support, and cable pull-down device. this last item being necessary for wells under pressure. Results obtained with this tool in the field are given. INTRODUCTION One of the key requirements for the introduction in the field of a new method of permanent well completion was the development of a new perforating method involving the use of a perforating gun small enough to go through tubing and with enough penetrating power to perforate the casing satisfactorily. The equipment described in this paper was basically designed to go through open-ended two-in. tubing and to perforate 5 1/2 in. casing. The equipment is illustrated in accompanying figures. NEW COMPLETION METHOD With the new permanent type well completion, the production string of casing is run and cemented through the potential producing zone. Open-ended tubing is -then run into the well and used to replace the drilling fluid with oil or water. The amount of the oil or water "cushion" can be varied according to the expected bottom hole pressure. Thus, if desired, the hydrostatic head can be greater, equal to, or less than the expected pressure within the zone to be perforated. Mud having been displaced, the tubing is -then positioned so that the end is above the zone to be perforated. The Christmas tree and flow lines are installed and tested. The drilling rig can now be removed from the well since it is not necessary for the perforating operation, thus resulting in a saving in rig tme. The perforating equipment is then moved to the well and the gun lowered through the tubing. After the gun passes out of the open end of the tubing, the depth measurements are correlated with casing collars and the gun placed opposite the zone to be perforated. Upon firing, the gun body is shattered and falls to the bottom of the well. Since the lubricator has been closed before firing, it is possible to withdraw the casing collar locator and head and then open the well immediately to the tanks. DESCRIPTION OF GUN An investigation of two basic types of Perforators, namely a bullet perforator and a shaped charge gun, was undertaken towards the development of a gun to meet the aforementioned requirements. Preliminary tests on various bullet perforator designs within the diameter limitations did not give results encouraging enough to warrant further study. Somewhat better results were obtained by using conventional shaped charges; but, as in the case of the bullet types, the necessarily small overall diameter proved to be a severe limitation and resulted in insufficient penetration. A shaped charge employing a new design was developed which gave excellent penetration. These shaped charges were designed to produce a large hole directed upward at a 45" angle. In order to obtain the maximum space possible for these charges, it was decided to use an expendable carrier made of drillable material, usually aluminum, which would shatter at the time of firing, leaving only harmless fragments at the bottom of the hole. The fixed spacing is five shots per ft. wit11

Jan 1, 1952

By N. M. Levine, W. M. Fassell

In this laboratory study the problem of aerative conditioning to separate chalcopyrite and pyrite from cobaltite was simply effected with a sulfy-drate collector and pH by proper choice of mixing variables. It was shown that for a given impeller type and geometry, selectivity is governed by a relation between the Froude number and a modified power number multiplied by impeller tank-diameter ratio wherein the power factor is expressed as power intensity or horsepower per unit volume of contained pulp. In applying this approach to other systems, other criteria would be used in place of selectivity. The purpose of this paper is to show how solutions to problems of aerative conditioning of pulps may be enhanced by development of mixing parameters in a given system through application of a fundamental approach to agitation. The term aerative conditioning refers to a combination of processes which include mechanical gas dispersion, gas solution in the liquid phase, transport of dissolved gases and other reactants to the reaction zone, physico-chemico interaction within the liquid phase and/or at the solid surface, and, finally transport of products away from the reaction zone. Important elements of these processes are: 1) Air-Liquid Contact Area: This area limits the rate at which a given gas may dissolve in a given liquid. Thus, an aerative conditioning system should be designed for optimum gas dispersion. 2) Diffusional Factors: Such factors as concentration gradient, diffusion film thickness, and diffusivity may govern the rate at which dissolved constituents are transferred to the zone of interaction at a particle surface and the rate at which reaction products are transported therefrom. Thus, such systems would be designed for maximum turbulence to effect good mixing and to minimize diffusion film thickness. 3) Solids-Turbulent Liquid Contact Time: Sufficient time should be provided for interactions to take place at solid-liquid interfaces. The region of maximum turbulence for mechanical systems being in the impeller zone, it is important that the number of passes through this zone be optimized. This means that a proper balance should be struck be- tween flow and turbulent power to achieve optimum conditioning time. Too high a flow rate and insufficient turbulence may be improper if the reactions are diffusion limited. On the other hand, too much turbulent power and too little time in the reaction zone due to insufficient flow power may also be limiting. 4) Geometry and Impeller Design: The design of the conditioning vessel and the type of impeller employed are important factors in determining the relative distribution between flow and turbulent power and the degree of gas dispersion. Maximization of power input would be to no avail if the distribution between flow and turbulent power were improper or gas dispersion insufficient. 5. Physical Chemistry: The reactions between dissolved gases and solid surfaces may be limiting if gas solubility, activation energies and/or free energies are not favorable. J. H. Rushton, E. W. Costich, and H. J. Everett' have summarized the mechanical elements of the combination of processes listed in a fundamental mixing equation. This equation was derived by dimensional analysis and relates the physical variables for a single impeller, centered in a cylindrical, vertical axis, flat-bottomed tank. It is, where T is tank diameter in feet; H is liquid depth in feet; C is height of impeller off tank bottom in feet; S is pitch of impeller; L is length of impeller blades; W is width of impeller blades; R is number of baffles; D is impeller diameter in feet; Np is the power number, pg/dN3D5; P is power in foot-pounds per second; g is the gravitational constant, feet per second; J is width of baffles; B is number of im-

Jan 1, 1961

By L. L. Handy

The most common method of identifying hydrocarbon-bearing strata in a well that penetrates many different formations involves measurement and interpretation of the electrical properties of the formations as determined by electrical logs. Even though this method is used extensively, and even though in a great many instances it is capable of indicating presence of oil or gas, situations arise for which it is extremely difficult, if not impossible, to deduce the presence of hydrocarbons. These situations may involve the following. 1. A thin formation, bounded by highly resistive formations, in which it is impossible to obtain the actual resistivity of the uninvaded zone with existing logging devices. 2. A formation in which invasion has been so extensive that a value for the uninvaded zone resistivity cannot be obtained. 3. A very shaly formation in which the resistivity index, I, is lower than that usually associated with productive formations. 4. Laminated formations comprised of thin productive sands separated by thin shale streaks in which the individual sand and shale streaks are too thin to permit measurement of uninvaded-zone resistivity with existing logging devices. 5. Productive formations in which the water saturation is high. To extend the utility of electric log interpretation to identification of hydrocarbons in all types of formations, there is strong incentive to find a method not subject to these limitations. Some time ago, in connection with research on the wettability of reservoir rock, an investigation was conducted in which the resistivities of cores were measured shortly after they were removed from a core barrel, and again after they had been extracted and restored to their original oil and brine saturation.' The resistivities after extraction were generally lower. Other tests made on the cores indicated that they were more nearly water wet after they were extracted; thus, it was assumed that the observed changes in resistivities were due to a change in wettability of the cores. Other experiments'," have shown that resistivities of rock samples are sharply dependent on wettability. These experiments have shown that oil-wet samples are more resistive than water-wet samples. To obtain an understanding of how the wetting properties of the surfaces of core material affect electrical resistivity, a series of experiments was conducted. Two groups of core samples were prepared for testing. One group contained brine, but no residual oil. The other group was saturated with brine, flooded with oil to a low water saturation, then flooded with brine to a final residual oil saturation. Resistivity measurements were made on each group. Both groups were then flooded with the original brine to which a chemical had been added that renders sand and clay surfaces preferentially oil wet, a so-called reverse-wetting agent. very little change in resistivity was observed in cores containing only water. The group containing residual oil, however, showed resistivity increases of 100 to 200 per cent. These experiments showed that the resistivity of a core containing oil could be altered by changing wet-tability of the core. Moreover, the possibility was introduced that reverse-wetting agents might be employed as the basis for a logging method for identification of oil-bearing strata. Since behavior of a porous rock containing gas and water might be expected to be similar to that of a rock containing oil and water, such a method should also be applicable to identification of gas-bearing zones. In principle the wettability of the invaded zone could be reversed without altering conductivity of the interstitial water or the hydrocarbon saturation therein. Those strata showing significantly increased invaded-zone resistivities would, therefore, contain hydrocarbons; those with no significant change would be filled only with water. Addition of a reverse-wetting agent to a hydro carbon -bearing zone which is, by nature, already preferentially oil wet would not result in an enhancement of its resistivity. It is generally believed, however, that most hydrocarbon-bearing strata are preferentially water wet.

By E. C. Tveter, L. A. Roe

The authors present a general outline of the theory and development of heavy-liquid application to mineral processing. Patent literature and processes are reviewed with special emphasis on liquid recovery systems which have been employed or proposed. Advantages and disadvantages of the process are discussed together with the recent developments which have revived interest in this old concentration method. The most important single factor in this resurgence of interest is concluded to be the narrowing gulf between chemical engineering and mineral dressing which has opened the field for new concepts of mineral plant design. A partial summary of patents on heavy liquid separations is included. In spite of the fact that heavy-liquid separation with organic liquids has been used in the laboratory for over 50 years, this process has never graduated to large scale commercial use for any extended period. However, a variation of this process, the sink-float or heavy-media separation process has found wide acceptance and is used to separate minerals from diamonds to gravel. Materials used to increase the specific gravity of the pulps used as heavy media include sand, clay, barite, magnetite, galena, hematite, atomized lead and ferrosilicon. Because of the greater ease of recovery, the magnetic materials, magnetite and ferrosilicon, are the preferred media today. In the U. S. alone, heavy-media iron ore plants with a capacity of over 10 million tons of concentrate per year are in existence. Heavy-media separation involving use of solid, inorganic particles suspended in water rapidly found a wide range of commercial use with the introduction of magnetic media. The minerals engineer is experienced in working with suspensions and quickly learned to develop and control such media at a cost compatible to the type of separation desired. The natural superiority of a heavy liquid with uniform chemical and physical properties has never been questioned since its first use in laboratory mineral separatory procedures. It offers the only method for gravity separation of fine particles of relatively close specific gravity. The most important early attempt to make heavy liquid separations commercial were made by the DuPont Co. which began experimental work on Virginia limonite in 1904. The appended "Partial Summary of Patents" compiled by W.L. O'Connell of The Dow Chemical Co. demonstrates the quantity and sequence of this work. Both inorganic and organic parting liquids were investigated and numerous patents were issued on the use and recovery of these liquids. The work culminated in the heavy liquid plant of the Weston Coal Co. at Shenandoah, Pa. Chemical engineering design of this plant was the responsibility of Francis I. and Hubert I. DuPont. Heavy-liquid separation of coal almost achieved commercial status at this plant which actually processed over 20,000 tons of coal. The liquids used were tetrabromethane, pentachloroethane and trichlo-roethylene, with liquid losses ranging from 8.9 to 12.4 oz per ton of cleaned coal. The reasons for failure of this plant are not clear but probably involved toxicity problems as well as other problems in chemical engineering. Excellent economics were reported. A more recent pilot plant was built in 1954-55 by Norris Goodwin for the Inerto Co. to treat hectorite clay. This plant employed carbon tetrachloride in jigs with liquid recovery by evaporation. Although good separation was achieved, incomplete removal of the CCl, from the clay prevented commercial operation. The only present operations known to the authors employing heavy liquids for gravity separations are limited to the use of calcium chloride in certain coal washers and bromochloromethane in a small batch operation (one ton solids per day) for the separation of beryllium metal particles from slag materials. This is not strictly a minerals beneficiation problem but it has demonstrated the feasibility of such separations. PROBLEMS Critics of early attempts to commercialize heavy-liquid separation of minerals summarize the draw-

Jan 1, 1963

By F. C. Bond

Since the Third Theory of Comminution was presented eight years ago (I) it has found increasing use in crushing and grinding problems. The practical utility of its wok index equation is quite generally acknowledged (2). However, its theoretical basis has been questioned in at least three technical articles (3) (4) ('). The purpose of this paper is to present experimental proof that it is scientifically correct. Particles under compressive stress are strained and deformed. They absorb strain energy, and when this locally exceeds the breaking strength, a crack tip forms. The surrounding strain energy flows to the crack tip, which rapidly extends and splits the rock, releasing the strain energy as heat. The initial energy flow causes additional crack tips in highly strained areas. If the compression is rapidly applied by impact, crack tips may form before the strain energy has reached equilibrium in the particle, thus decreasing the total work input required for breakage. The energy necessary to break is essentially the energy necessary to produce crack tips, since the energy necessary to extend the cracks to breakage is already present as strain energy in the deformed particles. After breakage nearly all of this energy appears as heat. The crack length cannot be measured directly. However, in particles of regular and similar shape the crack tip length is considered as equal to the crack depth, or crack extension necessary to break, so that the crack length equals the square root of one-half of the surface area. The Third Theory states that the useful work done in crushing and grinding is directly proportional to the total length of the new cracks formed. It can be confirmed by showing that a constant work input produces a constant length of new cracks when reducing the same material to different product sizes. This is done in the present paper on a wide variety of material. The constant work input was supplied by one revolution of the 12" x 12" laboratory ball mill used in making grindability tests by the Allis-Chalmers method (12) (13) The new crack lengths produced per mill revolution were measured from all available grindability test results at 28, 35, 48, 65, and 100 mesh on fifteen different ores, and were found to remain substantially constant for each ore at all mesh sizes. A new technique is used for the measurement of crack lengths. Size distribution plots of the mill feed and product are made by the Third Theory method (9) and the crack lengths are obtained from these plots by the procedure described in the present paper. The energy input required to produce a unit length is of fundamental importance in the size reduction of brittle solids. The crack length Cr is expressed in centimeters per cubic centimeter of solid material. It bears a definite relationship to the external surface area of the crushed or ground solid. A uniform particle shape must be assumed before the surface area and crack length can be evaluated. In this paper it is assumed that the relationship between the surface area and the particle volume of a particle d microns in diameter is the same as that of cube d microns on a side. The external surface areas of particles with a cubical breakage probably agree approximately with this rule, and correction factors can be applied when physical measurements of the surface areas are available for comparison. However, the assumption of equivalent cubes has been found satisfactory for most calculation purposes. Assuming equivalent cubes, one cubic centimeter of particles d microns in diameter will have a crack length Cr of v30.000/d centimeters, and a surface area of 60,000/d square centimeters. The specific crack length is thus equal to the square root of one-half the specific surface area. Where Sa is the surface area in square centimeters per gram and Sg is the specific gravity of the ground solid, then Cr = vSg . Sa/2 = 173.2/ vd (1)

Jan 1, 1961

By J. H. Swisher

When an Fe-Mn alloy is internally oxidized, the inclusions formed are MnO which contains some dissolzled FeO. In the internal oxidation reaction, not all of the manganese is oxidized; some remains in solid solution as a result of the high Mn-0 solubility product in iron. Taking these factors into consideration, the rate of internal oxidation of an Fe-1.0 pct Mn alloy is computed as a function of temperature, using available thermodynanzic data and recently published data for the solubility and diffusivity of oxygen in iron. The predicted and experimentally determined rates for the temperature range from 950 to 1350°C are in good agreement. ThE rates of internal oxidation of austenitic Fe-A1 and Fe-Si alloys have been studied extensively.1"4 Schenck et al. report the results of a few experiments with Fe-Mn alloys at 854" and 956C, and Bradford5 has studied the rate of internal oxidation of commercial alloys containing manganese in the temperature range from 677" to 899°C. When Fe-Mn alloys are internally oxidized, the inclusions formed are solutions of FeO in MnO, the composition depending on the experimental conditions. Since the thermodynamics of the Fe-Mn and FeO-MnO systems have been investigated,6"9 and since the solubility and diffusion coefficient of oxygen in y iron have been determined recently,' it is possible to predict the rate of internal oxidation from known data. The calculations used in predicting the rate of internal oxidation will first be outlined, then the results of the prediction will be compared with the experimental results of this investigation. PREDICTION OF PERMEABILITY FROM THERMODYNAMIC AND DIFFUSIVITY DATA Oxygen is provided for internal oxidation in these experiments by the dissociation of water vapor on the surface of the alloy. The dissociation reaction is: + H2(g) + [O] [1] where [0] denotes oxygen in solution. The equilibrium constant for this reaction is known as a function of temperature:' log As oxygen diffuses into the alloy, oxide inclusions are formed which are MnO with some FeO in solid solution. The reactions occurring are: [Mn] + [0] = (MnO) [31 and [Fe] + [0] = (FeO) [41 where [ Mn] is manganese dissolved in iron and (FeO) is iron oxide dissolved in MnO. The overall reactions may be written as follows: [Mn] + HOte) = (MnO) + H2(£) [5] and [Fe] + H20(g) = (FeO) + Hz(R) [61 The standard free-energy changes and equilibrium constants for Reactions [5] and [6] are known.6 Therefore the equilibrium constants for Reactions [3] and [4] may be obtained by combining known thermodynamic data for Reactions [I], [5], and [6]. For Reactions [3] and [4]: K = and For the present purpose, both the Fe-Mn7,8 and FeO-~n0' systems can be considered to be ideal, i.e., [amn] = [NM~] and (aFeO) = (NM~~) = 1 - (NFeO) where the Ns are mole fractions. These relations, together with Eqs. [I] and [8], permit us to compute both the oxide and metal compositions as a function of temperature and oxygen potential at any point in the specimen. For cases where the oxygen concentration gradient between the surface and the subscale-base metal interface is linear, the kinetics of internal oxidation is an application of Fick's first law: where dn/dt is the instantaneous flux of oxygen into the specimen, g-atom per sq cm sec; 6 is the instantaneous thickness of the subscale, cm; Do is the diffusion coefficient of oxygen in iron, sq cm per sec; p is density of iron, g per cu cm; h[%O] is the oxygen concentration difference between the surface and sub-scale-base metal interface, wt pct. B6hm and ~ahlweit" derived an exact solution to the diffusion equation for systems in which there is a stoichiometric oxide formed. They showed that the oxygen concentration gradient is given by a rather complex error function relation. For the Fe-Mn-0 system and for most other systems that have been studied, however, variations in oxide compositions are small and rates of internal oxidation are sufficiently slow that the deviation from linearity in the concentration gradient of oxygen is negligible. The mass of oxygen transported across a unit area of the specimen for the total time of the experiment is given by the mass balance equation:

Jan 1, 1969

By R. E. Boehler, N. C. Ludwig

At Buffington Station, Gary, Ind., Universal Atlas Cement operates fourteen 8 x 101/2 x 155-ft cement kilns in mill 6 and two 11 x 360-ft kilns in the Harbor plant. The No. 11 and 12 kilns in mill 6 are equipped with Manitowac recuperator sections. This report describes studies in measuring exterior shell temperatures on several of these kilns and the development of a traveling radiation pyrometer with certain novel features. Preliminary Work: At first various temperature-sensing devices were placed on the steel shell: 1) crayons with calibrated melting points, 2) colored paints with temperature-calibrated pigments, 3) aluminum paints with temperature-calibrated binders, and 4) metal-stem dial thermometers. The colored paints and aluminum paints failed to indicate the temperatures correctly. The crayons and thermometers did indicate fairly correct temperatures, but it proved impossible to apply enough of these on the shell to detect all the potential areas where hot spots might develop. Furthermore, considerable labor was required to apply these sensors and read the temperatures. Consequently no further work was done with these devices. Formation of Hot Spots: In the burning or clinker-ing zone of a cement kiln, the thickness of the protective coating and thickness of the brick govern the amount of heat transmitted to the steel kiln shell. Usually the protective coating consists of 4 to 8 in. of fused cement clinker. The formation of a hot spot is usually caused by loss of coating? that is, localized areas of the coating become thin or fall away from the refractory. This is generally caused by excessive temperature in the burning zone over a fairly long period of time. It may also be caused by a sudden thermal change in the burning zone. Variations in raw feed composition and in feed rate require changes in the fuel and air rates, and when these are not appropriately altered, conditions may develop in the kiln that will result in loss of coating. Luminescence on the kiln shell indicates that a hot spot has developed to a point that usually alters the refractory's thermal conductivity properties. When this thermal weakness zone occurs in the burning zone of the kiln, constant vigilance is required to protect it by maintaining proper coating. Even so, it has been the writers' experience that within a period of several days to about four weeks the hot spot usually recurs with greater severity. This necessitates shutting down the kiln and re-bricking the affected area. One of the prerequisites of a good burnerman is the ability to maintain a protective coating despite the many variables in operation. When he knows that it is getting thin or that an area has dropped off, he reduces the firing rate and kiln speed and brings feed into the affected area in an effort to rebuild the coating. But when powdered fuel is burned, the atmosphere of the kiln may prevent the burnerman's observing the condition of the coating closely at all times without taking off the fire. It is not considered good practice to do this frequently, as it imposes a thermal shock on the coating and upsets operation of the kiln. To help the burnerman scan the shell of the kiln along the burning zone, therefore, a radiation pyrometer, connected to a potentiometric recorder, was mounted on a slowly moving steel cable. The theory of operation, construction details, and adaptability of the radiation pyrometer are included in an excellent monograph' and also in a textbook.' Shell temperatures of the Atlas Cement kilns were measured with a Brown Instruments Div. low intermediate range Radiamatic unit, of range 200" to 1200°F, and a circular chart Electronik potentio-metric recorder, of range 500" to 1000°F. In Bulletin 59095M the supplier publishes standard calibration data (millivolts vs degrees Fahrenheit) for this radiation pyrometer, These data, however, apply only to flat surfaces having emissivities of unity. Calibration of Radiation Pyrometer for Use on Curved Surfaces: When applied to surface temperature measurements, the radiation pyrometer reading depends on the nature of the surface, the material of which it is composed, and also to some extent on the temperature of the surroundings. Although the present radiation pyrometer is designed to give a calibrated response under ideal (black body) conditions when used commercially, it must be calibrated empirically. The calibration procedure, given below, follows that described by Dike (Ref. 1, pp. 38-39). Calibration tests on plane and curved surfaces showed that the response of the radiation pyrometer was very

Jan 1, 1961

By G. F. Bolling

STUDIES of grain boundary migration in zone-refined metals have all shown that the rate of migration is greatly reduced by small added solute concentrations. However, it is apparent that a difference exists between boundary migration during normal grain growth and single boundaries migrating in a bicrystal to consume a substructure. To effect the same reduction in velocity in the two cases, much more solute is required for grain growth than for the single boundary experiments. One case is available for direct comparison; both Bolling and winegardl and Aust and utter' added silver and gold to zone-refined lead to study grain growth and single boundary migration, respectively. For comparable reductions in migration rates, about 500 times more solute was required to retard grain growth than to retard the single boundaries. A reason for this difference is suggested here. The rate of grain boundary migration is dependent on solute concentration and must therefore also depend on the solute distribution; i.e., regions of higher solute concentration encountered by a moving boundary must produce greater retardation and thus could determine any observed rate. A dislocation substructure can be the source of a nonuniform solute distribution since it can attract an excess concentration of certain solutes. In fact, it is probable that the solutes which impede grain boundary migration most would segregate most severely to a substructure for the same reasons. Thus a dislocation substructure present in a crystal being consumed could locally magnify the concentration of solute confronting an advancing grain boundary. In the single boundary experiments a low-angle substructure, within single crystals obtained by growth from the melt, was used to provide the driving force to move a grain boundary; in grain growth, no substructure of this magnitude was present. The increased solute concentration at subboundaries should be given approximately by C, = G e c,/kT, where t, is a binding energy and CO the bulk concentration. To account for the difference between the two experiments in the Pb-Ag and Pb-Au cases, C, must be the concentration impeding the single boundary migration, and a value of t, = 0.25 ev is necessary. This is reasonable, even though calculation on a purely elastic basis gives t, = 0.12 ev. because electronic effects must enter for silver and gold in lead. The compound AuPbz forms3 and the metastable compound AgrPb has been reported to nucleate at dislocations prior to the formation of the stable, silver-rich phase.4 Other observations support the hypothesis that a magnified solute concentration impedes the single boundary migration. For example, some crystals were grown by Aust and Rutter at concentrations of ~ 0.1 wt pct Sn and 2 x X at. pct Ag or Au which exhibited a cellular substructure, and in these crystals no boundary migration was observed. It is therefore evident that the higher concentrations at cell boundaries drastically inhibited migration. Inclusions would not have been responsible for this inhibition since according to recent work on cellular segregation,5 no second phase should have occurred in the segregated regions at the cell boundaries for the conditions of growth used, at least in the Pb-Sn system. In the purest lead, only the "special" boundaries observed by Aust and Rutter gave rise to the same activation energy as that obtained in grain growth. It is reasonable to suppose that the structure of special boundaries does not favor segregation at low concentrations and thus solute, or an inhomogeneity in its distribution, would have no effect. Random boundaries, on the other hand, are affected by solute and the substructure would enhance residual concentrations in the zone-refined lead, leading to a higher activation energy. It is clear, even without a detailed theory, that the apparent activation energies and exact solute dependence in the two experiments must be different as long as the non-uniform solute distribution produced by the substructure is important. Recrystallization experiments should also be susceptible to the same kind of local segregation at subboundaries or disloca tion cell walls; a suggestion similar to this has been made by Leslie et al.' Following the arguments presented here, the effects of a given solute concentration would be like those observed by Aust and Rutter if segregation occurred, and like those of grain growth otherwise. This work was partially supported by the Air Force Office of Scientific Research; Contract AF-49(638)-1029.

Jan 1, 1962

By Werner Schwartz, Wolfgang Haase

Lead sulfide concentrates, which may include other lead concentrates, are sintered on an up-draught sintering machine without the addition of any diluting agents or fluxes. Subsequently they are melted in an oil- or gas-fired rotary furnace. The sintering and melting processes are based upon the following roast-reaction: PbS + 2 PbO = 3 Pb + SO, PbS + PbSO, =2 Pb + 2 SO, For obtaining a lead bullion free from sulfur, the sintering process is carried out in such a way that the sinter product contains a small amount of excess oxygen above that to react with the sulfides. At the end of the melting process, when the reactions are finished, the remaining small amount of oxide residues is reduced with coal to which a certain percentage of soda ash (about 1 pct of the lead bullion) is added. For the lead smelting process described neither coke nor fluxes—except soda ash—are required. This process is being utilized by a European smelter successfully and with a high lead recovery. The consumption figures for the smelting of 100 tons per day of lead concentrates are indicated. The lead content of the lead concentrates from modern ore dressing plants ranges from 65 pct to above 80 pct. In most lead smelters of the world these concentrates are smelted in a blast furnace. For blast-furnace smelting the concentrates have to be desulfurized and agglomerated by sintering. A requirement for the perfect operation of a down-draught sintering machine and of a blast furnace is a maximum lead content in the feed of 40 to 45 pct. For this reason, some lead concentrates have to be diluted by adding return slags, limestone, and possibly iron oxide and sand. As an example, 100 tons of lead concentrate with 72 pct Pb would contain 13.5 tons of gangue (including the zinc). To produce a perfect sinter with 42 pct Pb it would be necessary to add 70 tons of flux and return slag, more than five times the original weight of the gangue, to the sinter mix and blast-furnace charge. A correspondingly large amount of coke would be required in order that all of these materials reach the heat of formation and the melting temperatures of the slag (1200" to 1400°C) inside the blast furnace. The roast-reaction process presents a possibility for lead recovery without dilution of the concentrates. In this process the concentrate mixed with coal is placed upon a Newnam-hearth and air is blown through nozzles into the heated mix. AS a result metalllic lead and a relatively great amount of so-called .'Grey Slag" with a lead content of 25 to 35 pct are formed. The slag is sintered to eliminate sulfur and, after addition of the requisite fluxes, treatt:d in a blast furnace. Owing to the poor recovery of lead from the hearths and to the unavoidable heavy hand-work plus the risk of poisoning this process is utilized in very few 112ad smelters today. Since in mxny countries of the world coke is expensive and difficult to obtain, it appeared feasible to use the principle of the roast-reaction by modern sintering and melting methods with recovery of the lead in electric, or oil, gas, or coal-fired furnaces. Two processes are utilized on an industrial scale: A) Lead smelting in the electric furnace of the Bolidens Gruv A/B in Sweden, as described by S. J. Walldcn, N. E. Lindvall, K.G. Gorling, and S. Lundquist. B) The self-fluxing lead smelting of Lurgi Gesell-schaft fiir Chemie und Huttenwesen m.b. H., Frankfurt a M, Germany, which is described in this paper. In the Boliden process referred to above the sinter mix is pelletized by enveloping return fines with layers of flue dust, limestone powder, and dried galena concentrate. The roasting and agglomeration are carried out on a down-draught machine, and a slight excess of sulfur is left in the sinter product. During the smelting in the electric furnance the roast-reactions occur and a slag poor in lead and a sulfur bearing lead are formed. This latter is subsequently oxidized in a converter to obtain lead bullion and dross. The Lurgi-process achieves the maximum possible extent of the roasting reaction on the sintering machine. The wet flotation concentrates are blended with return fines (lead content 70 to 80 pet), any existing flue dusts and lead slimes—but without the

Jan 1, 1962

By Reaves. M. J.

The title of my talk, "What Happened to the Uranium Boom?" is old news. Certainly it is for this group. All of us that make our living in uranium know that the boom of the last half of the 1970's is over. U.S. production has been exceeding consumption by more than two to one. Mines and mills are closing and yellowcake prices have been dropping for over 20 months. The gloomy outlook for the industry in the near term has been well documented by soothsayers of various descriptions, your daily newspapers, and in the Nuexco Monthly Reports. I'd like to attempt to describe the next upturn in the market (speculate, really) based upon the clues we're seeing now. In order to do that, I'd first like to go over briefly, some of the market factors that contributed to the recent price drop and resultant production cutbacks, and then hypothesize on the way these factors are changing and will change. Market prices are greatly affected (maybe even entirely determined) by buyer perceptions. This is particularly true with uranium, because of the long lead times associated with nuclear plant construction and also with conventional mine/mill development. Before the price rise (say, 1975) utility uranium buyers believed that: 1) U.S. producers would have difficulty expanding to meet U.S. demand. 2) Australian and Canadian production was essential to avoid shortages in the early 1980's. 3) Uranlum prices would continue to rise as demand exceeded supply. 4) Enrichment capacity would become inadequate. It was thought necessary, therefore, to build enriched inventory in the early 1980's for use in the late 1980's. Artificially accelerated expansion of the uranium producer industry was necessary to accommodate anticipated enrichment demand. Current perceptions are largely the opposite. These are the beliefs that were held most of this year and late last year as prices dropped. 1) U.S. production is far in excess of domestic need. Contraction of the U.S. production lndustry is necessary. 2) Canadian and Australian supply is optional and not essential. Producers in those countries are expanding mainly by displacing higher cost production and not because they fill a void, 3) Prices may be essentially stable for some time. 4) Enriched uranium is in excess supply. That is 1981. 1982 is shaping up to look like this: 1) Prices will have bottomed out. (That is not Nuexco's opinion necessarily, by the way, but it is my opinion.) 2) There will still be substantial utility inventories, but fewer spot sales. 3) Canadian and perhaps Australian sellers will have made substantial sales in the U.S. and will be aggressively seeking more. 4) U.S. production will have been dramatically curtailed. U.S. utilities that wish to con- tract long term will have difficulty in finding domestic sellers. Concern will develop about the availability of U.S. production capability. Virtually all long term con- tracts signed will be with non-U.S. sellers. 5) An awareness will begin to develop among U.S. buyers that we are approaching a period of dependence upon foreign uranium (which will be true). The history of the uranium market has been one of dramatic changes and overreaction to those changes. The rapid price rise of a few years ago generated excess U.S. production capacity and the rapid price drop of the last two years will almost certainly result in too little capacity. It will soon be difficult for U.S. buyers to buy domestic material except on the spot market. The question is, "will they care?" The lack of demand, of course, is the underlying reason for the current poor health of the uranium industry. In 1972, 1973 and 1974 collectively, there were 105 nuclear reactors ordered in the U.S. That ordering rate was expected to continue and accelerate throughout this century. In 1975, 1976, 1977, 1978, 1979, and 1980 altogether, there were 56 more reactors cancelled than ordered. The net growth of our only customer since 1974 has been a negative 56. TO put this in perspective, if these 56 reactors were operating now it would more than double present U.S. uranium consumption. Underlying lack of demand is something that is simply not going to change in this decade. Time is going to be required. The NRC indicates that the maximum feasible number of new reactors that can be licensed each year is six. That would increase uranium consumption by only 10% per year. New reactors, if ordered tomorrow, would not generate new uranium demand until after 1990. Even so, United States' consumption of uranium will rise from the 1980 level of 18 million pounds per year, to

Jan 1, 1982

By R. F. Shurtz

R. Duval (Mining Engineer, Ancien eleve de PEcole Polytechnique, Paris, France) I do not agree with the Eq. 3, reading: m =1/100- [(0.214x30.4) + (0.7B6 x0.00)] =6.5pct CaO If 0.214 and0.786 were proportions by weight, the equation would represent the well known mixtures law of the conventional arithmetics and 6. 5 pct CaO would be the correct average content. But it is not the case as the author states: "In samples consisting of single grains of mineral, those grains must, as already mentioned, be either of dolomite or magnesite. Since 78.6 pct of the deposit consists of magnesite and 21.4 pct of dolomite (excluding for present purpose the presence of other minerals), for any single grains picked at random the probability will be 0.214 that is it dolomite and0.786 that is it magnesite. In 1000 such samples the expected numbers of dolomite and mapesite grains will be 214 and 786 respectively." 0.214 and 0.786 would be proportions by weigbt under the necessary condition that all grains of dole mite and magnesite should have an identical weight. Obviously it is not the case, as the specific gravities are not the same for mapesite and dolomite and the volumes of the grains are different. Furthermore, because of these differences the conditions for a random sampling are not fulfilled and we are not authorized to state that the probabilities are, respectively, 0.214 and 0.786. The author however makes a simple application of Eq. 1: M = 1/n— ? fi x i . n Should we deduce that this relation is wrohg? Not at all, but when applying Eq. 1 you must not overlook what it actually. means. Eq. 1 gives a definition of the arithmetic mean of a total of n observed values Xi and nothing else. But the average conteht of a deposit has not the same significance. It is the ratio between the weight of concerned mineral in the deposit and the total weight of the deposit. As from 1000 particles the 214 of dolomite and the 786 of magnesite have not the same weight, the two definitions do not concur, and when applying Eq. 1 the result is an arithmetic mean of figures which has no connection with what is named average contentof a deposit. The situation is similar to the calculation of an average velocity. If a car travels a first mile over at 30 miles per hr and a second mile over at 60 miles per hr, when applying formula 1 you find as average velocity for the 2 miles: 30+60 ------- - 45 miles per hour. Many people calculate in this way and they do not realize that a mistake is involved. In fact the definition of he average velocity for the 2 miles is the quotient of the distance of 2 miles by the time (in hours) necessary for 2 miles travel, i.e.: 2 ---------- = 40 miles per hr. 1 + 1 30 60 In other words, the average volocity wanted is not the arithmetic but the harmonic average of the two velocities. The above mentioned bias in the calculation of the average contents of deposits is frequent, even in the assessments made by experienced engineers and is independant of what is named the sampling error. In order to supress the bias and to be able to use Eq. 1, you must apply a correction. An example on the subject can be found in an article by Duval et al. in the January 1955 issue of the ''Annales des Mines" (French), page 19. R. F. Schurtz (Author's Reply) Mr. Duval's position is quite correct. The proportions shown for dolomite and magnesite., respectively, of 0.214 and 0.786 are, in fact, proportions by weight uncorrected for specific gravity. In our day to day operation of producing magnesite from these mines at a very substantial rate, we do not normally make corrections for the difference between the specific gravity of dolomite and that of magnesite. If these corrections are made in Eq. 3 as shown in my article, then the numbers of grains turn out to be in proportions of 0.226 dolomite and 0.774 mapesite instead of the values actually shown in the equation. For the purposes of our work, and in view of the inherently low accuracy of the data, this correction was not deemed worthwhile making.

Jan 1, 1961

By L. H. Van Vlack, T. Y. Tien, R. J. Martin

The phase relationships of the MnTe-MnS system were studied by DTA procedures. There is an eutectic at 810°C with about 10 mole pct MnS-90 mole pct MnTe. An eutectoid occurs at about 710°C with approximately 7 mole pct MnS where the MnTe(NaCl) solid solution dissociates on cooling to MnTe(NiAs) and MnS. There is very little solid solubility of MnTe in MnS. ALTHOUGH MnS may exist in three different crystal forms,' only the NaC1-type phase is stable.2 Above 1040°C, MnTe also has the cubic NaC1-type structure. Below that temperature, MnTe changes to the NiAs-type structure.3 This phase transition is rapid for both heating and cooling. As a result the high-temperature crystal form of MnTe cannot be retained at room temperature. Because MnO, MnS, and MnSe are all stable with the NaC1-type structure, and MnTe has this structure at high temperatures,4 solid solution formation could be expected among these compounds. It is interesting to note, however, that a complete series of solid solutions exist only in the MnS-MnSe system,' and that the solid solution is quite limited in the MnO-MnS system.' The MnSe-MnTe system possesses a complete series of solid solutions at high temperatures with separation at lower temperatures.7 Although ion size may be critical in the miscibility of MnO-MnS, it is quite possible that the bond type plays a more important role with the miscibility of MnSe-MnTe. This would permit us to speculate that the miscibility gap would be extensive in the MnTe-MnS system. EXPERIMENTAL Preparation. The samples were prepared by mixing and compacting MnTe and MnS powders. The MnS was previously prepared through the sulfur reduction of Mnso4.8 The MnTe had been prepared by mixing and compacting double vacuum distilled metallic manganese and high-purity tellurium in stoichiometric ratio modified with 1 wt pct excess tellurium. The compacted powders were put in a graphite crucible which was sealed in an evacuated vycor tube. The free space in the vycor tube was made minimal to reduce the loss of tellurium. The sealed assembly was then heated slowly to about 500° C where the free manganese and tellurium reacted vigorously, melting the MnTe which formed. Only one phase, MnTe, was detected by X-ray powder patterns and metallographic techniques. Each compact of MnTe-MnS was placed in a graphite crucible and then sealed in an evacuated vycor tube. The samples were heated at 1250°C for 4 hr and furnace-cooled. Microscopic examination revealed no third phase beyond MnS and MnTe. A typical microstructure is presented in Fig. 1. Identification. X-ray powder patterns were obtained using 114.6 mm Debye-Scherrer camera and Fe-Ka radiation. Mixtures of cubic MnS and hexagonal MnTe were observed in all of the compositions prepared. No lattice parameter change was noticed among different compositions, indicating no solid solution could be retained at room temperatures between these two end-members. A lattice parameter of 5.244Å for MnS was obtained by the Nelson and Riley9 extrapolation method using the diffraction lines of (h2 + k2 + 12) equal 12, 16, 20, and 24. The values, a = 4.145Å and c = 6.708Å, for hexagonal MnTe were obtained from the (006) and (220) lines in the back-reflection region. These values agree well with the values reported by Taylor and Kag1e.10 Differential Thermal Analysis. A differential thermal analysis procedure was used to determine phase relationships since the high-temperature equilibrium conditions could not be retained for examination at room temperature, even when the sealed samples (~0.5 g) were quenched in water. The samples were sealed in an evacuated 4 mm vycor tube with a recess in the bottom to accept a thermocouple. An Al2O3 reference was similarly prepared and the two placed within a piece of insulating fire brick to dampen spurious temperature changes within the furnace. The furnace was controlled by a mechanically driven rheostat which increased the temperature at a rate of about 15°C per min. Known phase changes in the Pb-Sn system1' and the a-to-ß quartz inversion12 were used for calibration

Jan 1, 1969

By Lawrence A. Roe

The magnetic reflux classifier, which utilizes the combined effects of magnetic fields and a hindered settling classifier, is a new tool for determining the quantity and quality of middlings in fine-sized magnetite concentrates. Results are given for processing a typical taconite ore, and a sketch of the apparatus is included. IN examining magnetite ores and beneficiated products it often becomes necessary to make critical studies of the amount of grinding necessary to produce the desired degree of magnetite liberation. In the past this has been accomplished by laboratory heavy-liquid tests, which provide a method for selectively removing middling particles and free magnetite from various-sized fractions. Examination of the various products under the microscope results in fairly accurate determination of the degree of liberation. The method is quite efficient on sizes coarser than 325 mesh. Thus the heavy-liquid method of middling separation was satisfactory until the advent of present day magnetic taconite studies. When magnetite concentrates ranging from 70 to 100 pct —325 mesh are studied it becomes apparent that older methods of determining liberation size are not satisfactory and that there is need for a new method. For example, some of the low-grade magnetite ores of the Wisconsin and Michigan iron ranges require grinding to 100 pct —325 mesh to produce a magnetic concentrate containing less than 12 pct silica. Examination of concentrates from such ores often reveals that many of the middling particles consist of only very minor proportions of iron mineral. Thus it becomes important to be able to determine the degree of grinding necessary not only for complete liberation, but also for liberation of only 80, 85, or 90 pct of the total iron mineral content. Actually, complete liberation is never attained, but is often used to designate that degree of liberation necessary for production of high-grade concentrates. A rougher concentrate, produced after elimination of a coarse-sized tailing, can usually be subjected to a second grinding stage and concentrated into a higher grade product than could be produced from the same crude ore with one stage of grinding resulting in the same overall size reduction. This fact adds to the importance of being able to determine partial degrees of liberation on any magnetite ore. Standard laboratory methods such as heavy-liquid separation, microscopic grain counts, Davis tube magnetic separation, magnetic flocculation, classification, flotation, and others often are not applicable, or are prohibitive because of time requirements when large numbers of fine-sized magnetite samples are investigated. The Davis tube magnetic separator is an efficient tool to use in rejecting the non-magnetic mineral particles from an ore sample. The middlings discarded by the tube separator usually are so low in iron content that they can be considered relatively unimportant in liberation studies. This condition is caused by the extremely high flux density used in the Davis tube. This flux density ranges from four to eight times the flux density produced by most of the powerful commercial machines in use today. Thus the problem resolves itself into a search for a method of selectively removing middlings from Davis tube magnetic concentrates which will be both rapid and efficient. Those methods showing most promise in the development of a process for isolating middlings from extremely fine-sized magnetic concentrates were flotation and magnetic flocculation. The use of flotation to remove middlings from magnetic concentrates is reported in the literature.'.' The flotation process is effective in removing middlings from a magnetite concentrate, but physical entrapment of fine-sized free magnetite in the silica-bearing froth is an undesirable feature. The flotation method of removing middlings requires time, effort, and precise control of many variables, and does not meet the required degree of middling isolation. Magnetic Flocculation Magnetic flocculation has long been resorted to"-" in efforts to upgrade magnetite concentrates. One of the new magnetic taconite plants now under construction on the Mesabi Range includes magnetic flocculation in the flowsheet' as an accessory process to remove high-silica middlings and free silica which has been mechanically entrapped in magnetite flocs. The use of magnetic flocculation as a laboratory method of making precise separation of middlings was further investigated, since it offered a rapid, simple method of accomplishing the desired result. Magnetic flocculation involves the subjection of a magnetic concentrate to a strong magnetic field, passing the concentrate in a highly flocculated condition to a hydroseparator or other classifiers of various types, and removing free silica and middlings as overflow products. In an attempt to utilize simple

Jan 1, 1954

By P. Greenfield, C. C. Smith, A. M. Taylor

J. E. Harris (Berkeley Nucclear Laboratories, England)—Greenfield et al.11 attribute abrupt changes in slope of their log o/log i curves for heat-treated Mg/0.5 pet Zr alloy (zA) to &apos;atmosphere&apos; locking. It is proposed here that a more reasonable explanation of the apparent strengthening at low rates of strain can be based on precipitation either during the preanneal or during the creep tests. All the tests were carried out above 0.5 Tm where solute atmospheres are likely to be largely evaporated2 and can migrate sufficiently rapidly so as not to impose any &apos;drag&apos; on the moving dislocations. McLean3 has derived an expression for determining the temperature Tc above which, due to the high-migration rate of the atmospheres, Cottrell or Suzuki locking can play no part in determining creep strength. This expression, which holds for an applied shear stress of not greater than 5 X 107 dynes per sq cm is: Tc/Tm= 7/6.8 - log10? where i = secondary creep rate The values for T, corresponding to the maximum and minimum reported creep rates at each temperature have been calculated from the data of Greenfield et al. These are given in Table VII. All the test temperatures were above T,, the margin being greater for the higher temperatures and for the lower strain rates where the breaks in the log s/log ? curves occurred. Dorn and his collaborators14, 17 have studied systematically the effect of solute hardening on the creep properties of an A1/3.2 at. pet Mg alloy. In the temperature range where strain aging occurred in tensile tests, abnormally high-activation energies for secondary creep were obtained but at temperatures above 0.43 Tm, solute alloying did not have any effect on the creep parameters. Moreover, there have been no reports of any strain aging phenomenon during elevated temperature tensile tests with ZA material.18 Instead of the observed strengthening being due to atmosphere locking, it is now proposed that precipitates play an important role in enhancing the creep strength of the material. There are two possibilities—precipitation of zirconium hydride during the high-temperature preanneal and/or precipitation of the hydride or a-zirconium during creep. On the basis of the former the results can be interpreted in terms of a critical stress being necessary to force the dislocations through or over preexisting precipitates. From the latter, if the strengthening is due to pre- cipitation during the test then hardening should be associated with a critical strain rate. At low rates of strain, time is available during the tests for precipitation to occur either directly onto dislocations (thus pinning them) or generally throughout the matrix (which would impede dislocation movements). Examination of the data of Greenfield et al. suggests that both mechanisms may be operative since they observed precipitation during creep and also found that their alloys exhibited high-creep strength in the early stages of the low-stress tests, i.e., before creep-induced precipitation had time to occur. It is not easy to understand why they considered that precipitation of zirconium hydride is unlikely to occur at 600°C while it can take place in tests in air at as low a temperature as 200°C. Precipitation of the hydride during the preanneal cannot be ruled out merely on the basis of metallographic examination. Hydride precipitates in ZA type alloys are very small and can only be accurately resolved in the electron microscope.9 For example, in this laboratory20 hydride platelets with major dimensions <(1/10) µ have been observed by electron transmission through thin film specimens of hy-drogenated ZA material. Complex interactions between dislocations and such particles are illustrated in Fig. 12. Additional evidence for precipitation during pre-annealing is provided by the data presented in Greenfield&apos;s Fig. 1 and Table IT. These show that the creep strength at 200o and 400°C increases with the time of preanneal at 600°C. Such increases cannot be explained on the basis of increases in grain size alone for further improvements in strength were observed when the material was annealed for longer times than that required to stabilize the grains. Although the main discussion is confined to ZA material, similar arguments can be used against the strain aging hypothesis proposed to explain the binary Mg/Mn alloy data. In this case no precipitation is possible during the preanneal, but precipitation-hardening during creep can occur.

Jan 1, 1962

By R. D. Daniels, T. W. Wood

The tensile properties of columbium and Cb-H alloys containing up to 455 ppm H were studied as a function of temperature and strain rate. Hydrogen, introduced into columbium at elevated temperatures, using a thermal -equilibrium technique, embrittled columbium most severely at about —77°C. This elnbrittle ment occurred even at hydrogen concentrations of an order of 20 ppm. At higher temperatures, the hydrogen tolerance of columbium increased in relation to the increased solubility of hydrogen in tile metal. Below this temperature hydrogen tolerance, as determined by ductility and fracture stress, increased slightly. Strain rate had little effect on the tensile results for cross-head speeds over the range 0.002 to 2.0 in. per min. Strain aging during the tensile test appears to explain the ductility mininmum at —77°C. The apparent increase in hydrogen tolerance at lower temperatures is attributed to the low mobility of hyhogen. Experiments were performed in which samples were prestrained in tension at room temperature and tested to failure at —196°C. Results suggest that hydrogetz segregation to preformed crack nuclei can cause subsequent embrittlement even at temperatures where hydrogen mobility is too low to cause embrittlement in a normal tensile test. COLUMBIUM is an example of the class of bcc metals with ductile-brittle transition temperatures sensitive to the presence of interstitial atom contaminants. Hydrogen is one of these embrittling contaminants. The embrittling effect of hydrogen is less potent, perhaps, in columbium than in some of the other bcc refractory metals, but it is still a problem of both theoretical and practical interest. Unlike hydrogen in iron and steels, hydrogen in columbium is exothermically rather than endo-thermically occluded. The embrittlement process in exothermic systems has not been studied as extensively as that in endothermic systems, especially at hydrogen concentrations below the limit of solubility. The purpose of this investigation was to evaluate the embrittlement process in initially pure columbium as a function of hydrogen content, temperature, and strain rate. The Cb-H phase diagram, according to Albrecht et al.,1 is shown in Fig. 1. Columbium reacts exothermically with hydrogen producing a solid solution at concentrations of less than about 250 ppm (parts per million by weight) H at room temperature. At concentrations above the highly temperature-dependent solvus a second phase is formed. Like many similar hydrogen-metal systems,2 his system exhibits a miscibility gap with respect to hydrogen solution. Albrecht found the critical temperature of the miscibility gap to be about 140°C, the critical concentration to be 0.23 atom fraction hydrogen, and the critical pressure to be 0.01 mm Hg. Above 140°C there is a solid solution of increasing lattice constant extending across the phase diagram. Hydrogen concentrations of particular interest in this investigation were those below the limit of solubility in columbium. At hydrogen concentrations above the limit of solubility, columbium will contain the hydrogen-rich second phase and will be brittle under most testing conditions because the hydride generally precipitates as platelets with coincident matrix lattice strains.1&apos;3 At hydrogen concentrations below the limit of solubility, the tensile behavior of columbium is expected to be more sensitive to the interrelationships between hydrogen concentration and mobility and the testing variables such as temperature and strain rate. Literature references to the hydrogen embrittlement of metals, especially ferrous alloys and titanium alloys, are too voluminous to mention. It is only recently, however, that detailed studies of the hydrogen embrittlement of columbium have been undertaken. Wilcox et a1.4 studied the strain rate and temperature dependences of the low-temperature deformation behavior of fine-grained are-melted columbium (1 ppm H) and the effect of hydrogen content (1,9, and 30 ppm H) on the mechanical behavior of columbium at a series of temperatures for a single strain rate. A strain-aging peak was ob-served at about -50°C which was attributed to the presence of hydrogen in the metal. Eustice and carlson5 studied the effect of hydrogen on the ductility of V-Cb alloys at a series of temperatures over the range -196° to 25°C. Pure columbium was embrittled by 20 ppm H which produced a ductility transition at approximately -70°C. Ingram et al.6 studied the effect of oxygen and hydrogen on the tensile properties of columbium and tantalum. A minimum in the notched-to-unnotched tensile ratio of hydrogenated columbium was obtained at about -75°C, but because of the relatively large hydrogen content employed (200 and 390 ppm) the ductility

Jan 1, 1965

By J. A. Knox, R. M. Lasater, J. M. Tinsley

Chemical squeeze treatments have been used to provide temporary relief from certain production problems. The chemical fracture-squeeze technique, combining the effects of a fracturing treatment and a squeeze operation, has been more successful than conventional squeeze operations. Knowledge derived from well stimulation and reservoir engineering research provides a means for predicting the theoretical effective life of such a treatment. Analysis of theoretical equations and concepts developed allows selection of improved treatment techniques based on specific formation conditioins. Theory used in this analysis was developed as an extension of previous electrical model studies made to establish the expected flow and pressure profiles adjacent to a fracture system. The chemical fracture squeeze technique can be utilized in the economic application of corrosion inhibitors, emulsion breakers and paraffin and scale inhibitors. Application of this technique is shown to be effective. The slow return rate of injected chemicals, controlled by the resultant flow profiles and treatment variables, permits extended periods of chemical effectiveness. Results of field treatments are given, showing that the concepts outlined above for chemical fracture-squeeze treatments are valid and that applying this technique can help alleviate many current production problems. INTRODUCTION Much progress has been made in the last 10 to 15 years in developing chemicals for use in stimulating wells, maintaining production and protecting well equipment from damage due to corrosion. Not too many years ago, some wells seemed to dry up or wear out. In many cases the wells were produced as long as possible without any attempt at maintaining productivity. Even with the advent of new and better stimulation techniques, a rapid decline in production was observed. Methods of removing and, in some instances, preventing damage have been developed. Among thosc factors responsible for uneconomical production are scale, paraffin, corrosion, bacteria, water blocks and emulsions. Soluble scale-prevention chemicals have been developed1,2 that can be placed in a formation along with frac- turing sand. As the water produces back across this bed, the solid material dissolves slowly and can provide long-term protection from scale. However, bottom-hole temperature and salinity of produced water vary widely and both these factors influence the rate of solubility. Scale inhibitor composition is also a controlling factor. Some of the solid material may be crushed, increasing the surface area exposed to water and increasing the rate at which it dissolves. Some of the material may never be contacted by water and can be lost. However, this type of treatment has been very successful in many instances and has helped maintain economical production for extended periods of time. Liquid scale inhibitors, which are more widely applicable and more stable, have been developed in recent years; however, because they are liquids, their use has been restricted to treatment down the annulus, using metering pumps to provide proper concentrations in the produced fluid. This has prevented use in wells containing packers, in dually completed wells and in gas-lift and flowing wells. Wells that operate with an open annulus may also experience severe corrosion problems due to introduction of oxygen. Paraffin inhibitors3 have been developed that can be fractured into a well as particulate solids to be slowly dissolved in the produced fluid. These materials are not usually effective in wells with a bottom-hole temperature in excess of 120F since solubility rate may be too fast if that temperature is exceeded or if aromatic content of the oils is unusually high. Corrosion inhibitors have been developed that can be fractured&apos; into a well for long-term feedback, but development of a material with proper solubility or feed rate has been difficult. Corrosion inhibitors are available in many different forms. Liquids have been lubricated down the annulus or sticks or pellets dropped down tubing. Inhibitor squeeze treatments5 devcloped a few years ago led to development of inhibitors with particularly strong film-forming properties.6,7 This technique basically involves displacing a highly concentrated solution of the inhibitor into the formation through the tubing. Kerver and Hanson8 studied the adsorption properties of inhibitors on various types of formations. They showed that, even though the inhibitor was displaced radially into the true permeability, it could be produced back for a long period of time because of slow desorption from the rock. Methods developed for monitoring the return of these inhibitors generally have established 1 to 6 months as the effective limit before retreatment is necessary.9 Inhibitors displaced into the interstices of the formation sometimes cause emulsions that either hamper production or cause treating problems on the surface.

By R. H. Walpole, J. M. Kane

IN all probability the mining and metallurgical industry as a whole can demonstrate a larger ecorlomic return from installation of dust-control equipment than any other major industrial group. This fact has partially accounted for the marked increase of dust-control installations made during the past decade. While the primary objectives for installation of dust-collecting systems are improved working and operating conditions for men and equipment, the fact that an economic return can be anticipated on salvageable materials is an added advantage which shows in partial or complete equipment write-off. The conditions apply to most phases of the mining, milling, and smelting industry, both non-metallic and metallic. As with any mechanical devices, selection of suitable dust collector equipment involves evaluation of available products with characteristics most nearly meeting conditions of the application at hand. When there is valuable product to be collected, and/or when there are possibilities of air pollution or public nuisance, collector selection is often guided by the maxim of "highest available collection efficiency at reasonable cost and reasonable maintenance." A brief review of dust collector designs will permit outlining of major characteristics of each group. Final selection will involve detailed data against a background of the problem under consideration. The dry centrifugal collectors, see Fig. 1, represent a group of low cost units with minimum maintenance. They are subject to abrasion under heavy abrasive dust loads and to plugging with moist materials. Efficiency drops off rapidly on particle sizes below the 10 to 20 micron group. Because of the large amounts of —10 micron particles in most mining dust problems, they will normally be used as primary collectors and will be followed by high efficiency units. This combination is cspecially popular where the bulk of material is desired in a dry state with wet collection indicated for the final cleanup portion. In remote plant locations, dry centrifugal~ can be used alone if product in dust form has no value or if dust loading is light enough to eliminate a nuisance in the plant area. Where high efficiency dust colleotion equipment must be selected, choice will normally involve fabric arresters, wet collectors, or high voltage Electro-Static precip-itators. Fabric arresters, see Fig. 2, rely on the passing of dust-laden air at low velocity through filter fabric. Velocity ranges from 1 to 3 fpm for the usual installation and may be as high as 10 to 20 fpm in arrangements where automatic frequent vibration or continuous cleaning of the filter media is employed. Fabric is normally suspended in either stocking type or in an enlvelope shape. Collection efficiency is excellent even on sub-micron particle sizes. Equipment is bulky, must be vibrated to remove the collected dust load, and is restricted in applications from temperature and moisture standpoints. Condensation of moisture on the fabric filter mcdia causes plugging of the passages with great reduction in air flow. Temperatures for the usual medias of cotton or wool are 180" and 200°F maximum, although the introduction of synthetic materials such as nylon, orlon, and glass cloth have increased the possibilities of this type of collector for higher temperature applications. The wet-type collector may employ a number of different principles so that entering dust particles in the gas stream are wetted and removed. Principles usually include impingement on collector surface or water droplets, often in combination with centrifugal forces. Variety of wet collector designs is indicated by typical collectors illustrated in Figs. 3 and 4. Collection efficiency is a function of the particular design, although the better collectors will have high collection efficiency on particles in the 1-micron range. Wet collectors have the advantage of handling hot or moist gases, take up small space, and eliminate secondary dust problems during the disposal of the material. At times collection of the material wet is a disadvantage. Wet collectors may also be subject to corrosion and freezing factors. The high voltage Electro-Static precipitator, see Fig. 5, is probably the most expensive type of high efficiency collector. It finds its applications generally in problems in which collectors previously discussed cannot be employed. Its collection efficiency is based on its design features and can be excellent on the finest of fume particles. Material is normally collected dry. Gas temperatures are of no great concern as long as condensation does not occur within the dry type of precipitator and the temperatures do not exceed the limits for materials used in its construction. As with the fabric arrester, provisions

Jan 1, 1954

By M. D. Arnold, P. B. Crawford, P. C. Hall

An experimental study has been made to determine the optimum flooding pressures for four different oils. The oil formation volume factors ranged from 1.08 to 2.13, and solution gas-oil ratios ranged from about 200 cu ft/bbl to 2,250 cu ft/bbl. Viscosities ranged from 0.38 to 0.95 cp at the respective bubble points of the fluids and from 0.7 to 20 cp at atmospheric pressure. Water floods were conducted at various pressure levels from run to run. The recovery as a function of flooding pressure was found to be different for each fluid, with optimum gas saturations ranging from 7 up to 35 per cent. The data indicate that substantially higher recoveries may be obtained if water floods are conducted at an optimum pressure and that this optimum pressure is a function of fluid properties. The same core was used for all tests, and the reproduction of saturations for various runs indicates that wettability in the predominantly water-wet core did not change. INTRODUCTION A paper was presented by Bass and Crawford&apos; which described an experimental study of the effects of flooding pressure and rate on oil recovery by water flooding. This work was conducted using high-pressure models operated in a manner similar to that of an actual reservoir, with gas saturations being obtained by a solution-gas-drive mechanism. They found that the greatest oil recovery was obtained for the system studied by flooding in the presence of a 5 to 7 per cent gas saturation. Another experimental study simulating field conditions was presented by Richardson and Perkins.&apos; They used an unconsolidated sand pack containing kerosene-natural gas fluid and interstitial water. They flooded at various pressures and flooding rates. For their system it was found that the recovery was independent of the pressure level at which the water flood was performed. Kyte, et al," found that oil recovery by water flooding was increased as the free gas saturation at waterflood initiation was increased. However, after the initial gas saturation was increased above 15 per cent, the increase in oil recovery tended to level off. All of their runs were made at the same pressure using a light oil saturated with helium. The desired gas saturation was obtained by injecting helium into the core. Dyes&apos; made calculations which showed that an optimum gas saturation of 12 to 14 per cent may result in an increase in oil recovery of 7 to 12 per cent over that obtained by flooding at the bubble-point pressure. Others have also found that the presence of a free gas saturation may increase the waterflood oil recovery. In each case shrinkage was small and changes in fluid properties with respect to pressure were small. A careful review of the literature reveals that at the present time there is a wide difference of opinion on the factors affecting waterflood recoveries. This diversity of opinion is probably due to the fact that very little research has been done which has taken into account the many variables existing in an actual field being water flooded. Since the literature contains little information on high-pressure waterflooding studies using various types of reservoir fluids, it was believed appropriate that such a study should be made. EQUIPMENT AND PROCEDURE All tests were made using the same consolidated sandstone core. Torpedo sandstone was used to turn a cylindrical core 13.5-in. long and with a 2.92-in. average diameter. The core had a porosity of 28 per cent and a permeability to brine of 146 md. This brine was made up by adding 20,000-ppm sodium chloride and 30,000-ppm sodium nitrite to distilled water. This was used as connate water and flooding water. No fresh water was ever brought in contact with the core, as tests showed the sandstone contained argillaceous material which swelled in the presence of fresh water and plugged the stone. The core was sealed in a section of 6-in. N-80 tubing with Woods metal filling the annulus. The core was mounted horizontally; an injection well was placed in the center of one end and a production well in the center of the other. Pressure control was maintained by placing a back-pressure regulator (upstream control) on the producing well. The "live" oil was stored in a separate bottle and water was injected into this bottle to displace the oil for saturating the core using a two-cylinder standard-proportioning pump. This same pump was used for water flooding the core at a constant rate. This system was enclosed in water jackets and the temperature was automatically main-

By Eckhard Nembach

The phase diagram of the system Pb-In has been investigated between -27° and + 40°C, using nzainly X-ray dijfraction. In accordance with t her mo dynamic measurements by Heumann and Predel, a segregation occurs at low temperatures, though not in the form of a nziscibility gap. THE phase diagram of the system Pb-In has been the subject of extensive investigations,1&apos;1 but recently Heumann and prede13 concluded from their thermodynamic data that a new feature should occur below room temperature. These authors observed that the maximum values for the enthalpy and entropy of mixing, which occurred at a composition of 50 at. pct Pb, were +400 and —1.7 cal per g-atom deg, respectively. From this the authors estimated that a miscibility gap should occur below 30°C, centered at 50 at. pct Pb. Resistivity measurements seemed to support this view. These authors proposed the phase diagram outlined in Fig. 1. Three phases exist at 30°C: the tetragonal indium phase with c/a > 1, the tetragonal intermediate phase a, with c/a < 1, and the fcc lead phase. During an investigation of the superconducting properties of Pb-In alloys. it has been observed4 that aging a specimen with 50 at. pct Pb for 14 days at -18°C decreased the superconducting transition temperature about 0.13"K and tripled the transition width. In this paper, the results of an investigation of the Pb-In phase diagram in the temperature range from — 2T to +40°C are reported. Superconductivity and X-ray methods have been used. 1) SPECIMEN PREPARATION The materials were provided by the American Smelting and Refining Co. According to the manufacturer their purity was 99.999 pct. The weighed amounts of the constituents were sealed in quartz tubes under an atmosphere of 10 torr helium, mixed for 24 hr in a rocking furnace at 380°C, quenched in ice water, and homogenized at 20" to 30°C below the solidus line, established by Heumann and Predel. The annealing times were 144 hr for specimens containing Less than 30 at. pct Pb and 36 hr for the remainder. 2) SUPERCONDUCTIVITY EXPERIMENTS The specimens were quenched from the homogeniza-tion treatment into ice water and their superconducting transition temperatures T, measured. The procedure used has been described in Ref. 4. The transition was detected by the change of the mutual induc- tance of two coaxial coils containing the sample. T, was defined as the temperature at which 50 pct of the total change in inductance had occurred. The repro-ducibility with which T, could be measured was i0.002"K. Then the specimens with lead contents between 38 and 75 at. pct were aged for 7 days at temperatures between -30" and 40°C. If this treatment caused T, to change by more than 0.005"K or the width of the transition to increase by more than 0.002"K, it was concluded that the specimen had undergone a phase change and no longer consisted only of the fcc lead phase: as it did immediately after homogenizing. The result is shown in Fig. 2. From this one can estimate at what temperatures and concentrations phase changes occur. The X-ray measurements were based on these preliminary results. 3) X-RAY EXPERIMENTS Because of the softness of the material, relatively coarse powders. 75 p, had to be used, which were filed in a helium atmosphere from homogenized specimens. The powders were annealed at least 30 min at temperatures between 120" and 16OJC, depending on their concentration, and quenched in ice water. Then their X-ray patterns were taken at -178°C with a Picker diffractometer, model 3488K, and a cold stage. on which the specimen was in thermal contact with a liquid-nitrogen reservoir. In this way the following relation was established for the fcc lead phase: a = 4.697 + 0.00247C for 40 5 C 5 75 11 where n is the lattice constant (A) and C is the at. pct of lead. The coarseness of the powder made it impossible to use lines with 0 > 75 deg; therefore n was averaged from lines with 45 deg 5 0 5 75 deg. The results were reproducible to within i0.05 pct. Relation [I] is very similar to the one found by Heumann and Predel at room temperature. Following this, homogenized specimens with compositions between 15 and 56 at. pct Pb were aged for at least 10 days at temperatures between -27" and

Jan 1, 1970