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By C. E. Lundin

The Ce-H system was investigated in the temperature range, 573° to 1023°K, and the pressure range, 10-3 to 630 Torr, as a function of 'composition up to 72 at. pct H. Families of isothermal arid isopleth curves were plotted from the pressure-terr~perature-composition relationships. From these curves the solubility relationships were determined for the system. The isopleths are analytically represented by equilibrium dissociation pressure equations. The relative partial molal enthalpzes and entropies of solution of hydvogen in the systerrz were calculated fronz the dissociation pressure equulions and are tabulated. The integral free energies, enthalpies, and entropies of mixing in the Ce-H system were determined from the relative partial quantities and are also tabulated. The standard free energy, enthelpy, and entvopy of reaction of the dihydride phase at kcal per kcal per mole H2, and ?S° = -34. 1 cal per deg mole H2, respectively. The equilibrium dissociation pressure equation in the two-phase region is: UNTIL recently very little was known of the detailed solubility and thermodynamic relationships of the Ce-H system. Two previous investigations1,2 are noteworthy. However, significant discrepancies and omissions exist on analyzing them. The work of Mulford and Holley1 on cerium did not clearly delineate the boundaries of the two-phase region, Cess - CeH2-x. The plateau partial pressures were not thoroughly defined and were considerably displaced in pressure compared to those from the work of Warf and Korst.2 These latter authors concentrated their studies primarily from 823° to 1023°K in the pressure range of 1 to 760 Torr. No data were determined to outline the regions of primary solid solubility and the hydride phase. Also the establishment of the plateau partial pressures was rather limited in scope. In neither work was a treatment conducted of the relative partial molal enthalpies and entropies of solution of hydrogen in the single-phase regions and the integral thermodynamic quantities of mixing throughout the system. Therefore, it was the objective of this research to determine the complete equilibrium solubility relationships and thermodynamic data for the system by pressure-temperature-composition studies. EXPERIMENTAL PROCEDURE The cerium metal for this study was donated by the Reno Metallurgy Research Center of the Bureau of Mines. Total impurity content was 0.13 pct with only 60 ppm O. The metal was checked metallographically and contained only minor amounts of second phase compared to cerium from other sources. Specimen preparation was done in a dry box flushed with argon gas. The surface of a small rectangular piece of cerium (about 0.2 g) was filed with a clean, mill file. Final weighing was done in a tared enclosed vial containing argon gas. The specimen was then loaded quickly into the reaction chamber which was purged several times with high-purity hydrogen gas and then allowed to pump to about 10-6 Torr. The furnace was heated to the reaction temperature and the run started. The equipment used to conduct the hydriding was a Sievert's-type apparatus. Basically it consisted of a source for pure hydrogen, a precision gas-measuring burette, a heated reaction chamber, a McLeod gage, and a mercury manometer. Pure hydrogen was supplied by the thermal decomposition of uranium hydride. The 100-ml precision gas burette was graduated to 0.1-ml divisions and was used to measure the quantity of gas and admit it to the chamber. The reaction chamber was a quartz tube. Prior to each run, the cerium specimen was wrapped in a tungsten foil capsule to prevent reaction of the cerium with the quartz. Control of the temperature was achieved within ±1°K. Pressures in the manometer range were measured to ±0.5 Torr and in the McLeod range (10-3 to 5 Torr) to ±3 pct. The compositions of hydrogen in cerium were calculated in terms of hydrogen to cerium atomic ratio. These compositions were estimated to be ±0.01 H/Ce ratio. The technique used to study the equilibrium pressure-temperature-composition relationships of the Ce-H system was to develop experimentally a family of isothermal curves of composition vs pressure. The range of pressure through which each isotherm was developed was from 10-9 to about 630 Torr in the temperature interval, 573° to 1023°K. RESULTS AND DISCUSSION The hydriding characteristics of cerium are iso-morphous with those of the elements of the light-rare-earth group (lanthanum, cerium, praseodymium, and neodymium) wherein the region from the dihydride to trihydride is continuously single phase.' The structure of this phase is fcc.3 The heavy rare earths form a trihydride,2 which is hcp, separated by a two-phase region from the fcc dihydride phase. The Ce-H system is represented by the family of experimental isotherms in Fig. 1. Due to the small scale required to draw the curves, the experimental points are omitted; however, a total of 240 experimental data points were taken to prepare these curves. The solubility relationships can be deduced therefrom. Three distinct regions of partial pressure and composition can be seen. The region of cerium solid solution is represented by the rapidly rising isotherms in the dilute composition range. In accordance with Gibbs Phase Rule only one solid phase, the cerium solid so-

Jan 1, 1967

By C. B. E. Douglas

The following analysis was stimulated by a previous article on evaluation of the water problem at Eureka, Nev., which describes a method using formulas especially devised to calculate flow potential of extensive aquifers characterized by relatively even porosity and permeability throughout. The present discussion submits that the method was unsuitable for solving the kind of problem occurring at Eureka, where the amount of water available, rather than the flow potential, may have been the vital factor. IN an interesting article on evaluation of the water problem at Eureka, Nev.,1 W. T. Stuart describes how a difficult water problem, or one phase of it, may be evaluated by means of a small scale test. Test data are plotted by a method rendering, under certain conditions, a straight-line graph that can be projected to show how much the water table will be lowered by pumping at any specified rate for a given time. A formula is then used to determine the size of opening, or extent of workings, necessary to provide sufficient inflow to enable pumping to be maintained at that rate. At first glance this might seem the answer to a miner's prayer, but a word of caution is in order. It may not be the whole answer. Moreover, results obtained by the method described are reliable only for conditions approximating those assumed. Even where conditions do not meet this requirement, however, it may be possible to draw helpful inferences from the results, perhaps enough to facilitate another approach to evaluation of a problem. The two formulas Mr. Stuart used, the Theis formula and the one developed from it by Cooper and Jacob, were given field checks a number of years ago in valley alluvials by the Water Supply Div. of the U. S. Geological Survey and found to be reliable when the aquifer is very large in horizontal extent and sufficiently isotropic for the test well and observation wells to be in material of the same average permeability as the saturated part of the aquifer as a whole." Extensive valley alluvials, sands, and gravels can be evaluated in this way, and there are even cases in which the method could apply to porous limestones, such as flat beds of very large areal extent that have been submerged below the water table after extensive weathering. These are sometimes prolific sources of water for towns and industries. It is necessary for them to have been above the water table for some geologically long period of time in a fairly humid climate before submergence because the necessary high porosity and permeability, and large reservoir capacity, are the result of weathering, that is, of solution by the carbonic acid (H,CO3) in rainwater formed by the absorption of CO, from the air by raindrops, and this dissolving action must cease when all the H2CO3 has been consumed by re- action with the carbonate to form the more soluble bicarbonate. Consequently this weathering process is largely restricted to a zone that does not extend much below the water level, and submergence is necessary after the weathering to provide large reservoir capacity and good hydraulic continuity. On the other hand, water courses tend to form along faults and fractures in limestone, and to become enlarged by solution, well below water level when, as often happens, fresh meteoric water is circulated rapidly through them to considerable depth by hydrostatic pressure, as through an inverted syphon. Although the reservoir capacity of such water courses is relatively small they may extend far enough to tap more prolific sources. Cavities, and sometimes caves of considerable size, are found in limestones where the acid formed by the oxidation of sulphides has attacked them. This action can take place as deep below water level as surface water is carried by syphonic or artesian circulation, because the oxygen it carries in solution will not be consumed until it reacts with some reducing agent, such as a sulphide. Moreover, the formation of acid and solution of limestone in this way is not confined to the immediate vicinity of the sulphide. Oxidation of pyrite, for example, results in formation of acid in several successive stages, each taking place as more oxygen becomes available, as by the accession of fresh water into the circulation at some place beyond the sulphides. When the acid thus formed attacks the limestone, CO, is liberated and the ultimate effect of the complete oxidation of one unit of pyrite will be the removal of six times its volume of limestone as the sulphate and bicarbonate, both of which are relatively soluble. The reaction may be continued or renewed along a water course far from the site of the sulphides, where the small electric potential produced by contact with the limestone helped to start the reaction. Mr. Stuart refers2 0 caves in the old mining area in the block of Eldorado limestone southwest of the Ruby Hill fault at Eureka, Nev., and to the cavities encountered in drillholes in the downthrown block on the other side of the fault. Although he interprets these cavities as evidence that this formation was sufficiently isotropic (evenly porous and permeable) to give reliable results by the method he describes, they may, in fact, be entirely local conditions. There is reason to think they were probably formed

Jan 1, 1956

By O. J. Hassel, W. T. Maidens, J. H. Calbeck

The rotary gas fired reheating furnace used by the American Zinc Oxide Co. at Columbus, Ohio for Therotarygasfiredreheatingfurnacerefining lead-free zinc oxide is described. The outstanding features of this operation are that the color of the zinc oxide is greatly improved, sulphur is eliminated, and cadmium arethatrecovered without densifying the product to an objectionable degree. IN 1919 Leland S. Wemple obtained a patent for a process of reheating zinc oxide wherein the "coarsening of grain due to excessive heating was avoided." He taught in his specification that if solid carbonaceous material, such as lamp black, was added to the zinc oxide in proper amounts prior to reheating, objectionable sulphur compounds could be removed and the color would accordingly be improved and no objectionable densification would occur because of the relatively low temperature required. The situation that made this invention imperative was the newly opened zinc oxide plant of the American Zinc, Lead & Smelting Co. in Hills-boro, Ill. This was one of the early Western Type American Process zinc oxide operations. Characteristic of all of these early Western operations using Tri-State and Western ores was the great difficulty encountered in obtaining a product low enough in sulphur to compete with the Eastern Type American Process zinc oxides which were made from ores containing very low sulphur percentages. Wemple demonstrated that the refining process of his invention produced a superior color and although this was true and a most welcome feature, the primary purpose of the early refining operations at Hillsboro was to reduce substantially the high sulphur content of the crude zinc oxide. Although many and varied attempts had been made for refining zinc oxide none of the processes had a commercial history of any consequence until Wemple's invention became standard practice for the American Zinc, Lead & Smelting Co. in 1919 and their operations have been unique in that substantially all of their lead-free zinc oxide has been reheated since the first installation at Hillsboro. This process has become known in the industry as refining. The furnace developed by Wemple and continued in use by the company from 1919 until 1943 was unusual and merits some consideration by way of review in this paper. The furnace was essentially a double hearth coal-fired muffle furnace with a mechanical raking system consisting of a central shaft supporting six rabble arms in each muffle. The untreated or "crude" zinc oxide was fed onto the outer rim of the top muffle, moved to the center where it dropped to the lower muffle and progressed to the outer rim where it was discharged into an alloy screw conveyor. The retention in this furnace was extremely short, about 5 min, and the shallow zinc oxide bed on the hearths of the muffles was being continuously turned by the fast moving rabbles. Soft coal was burned on the grates below the lower muffle and the long yellow flame necessary to carry the heat around both muffles resulted in a very inefficient combustion of the fuel. The temperature of the top of the lower muffle seldom exceeded 65 °C although the oxide itself often reached 700°C before discharge. The capacity of this furnace was approximately 1/2 ton per hr. In our plant at Columbus it was necessary to keep four of these furnaces running in parallel to take care of the production because, as mentioned above, every pound of zinc oxide produced during these 24 yr passed through one of these refining furnaces. An essential part of this refining operation was the use of carbonaceous material admixed with the zinc oxide fed to the furnaces. Between 1 and 2 pct of a bran produced in the processing of cotton seed was added to all zinc oxide charged to the furnaces. The bran ignited on the top hearth and was still burning when the charge fell from the top hearth to the bottom hearth making a cascade of sparks. The rapid turning of the zinc oxide caused these particles of bran to flash on the hearths behind each rabble; but the combustion, of necessity, had to be complete by the time the charge reached the outer rim of the bottom hearth, otherwise the finished product would be contaminated with the charred particles of bran which would give the zinc oxide an unsatisfactory color. Although this operation was initiated to reduce objectionable sulphur percentages, as time went on new properties of the product were appreciated which made advisable continuing the refining process long after other methods of sulphur reduction became known in the industry. The particle size and particle size distribution, the absence of colloidal fines and perhaps a unique surface condition gave this product an outstanding performance when used in paints. The Wemple furnaces installed in Columbus in 1919 had to be rebuilt frequently and were extravagant in the use of fuel. The raking mechanism and the muffles required excessive maintenance expense and as the furnaces wore out the problem arose whether to continue along this line or to explore the possibilities of obtaining similar or better results in the simpler and more commonly used rotary furnace. To this end special research was initiated in 1941 on a small laboratory rotary

Jan 1, 1951

By J. B. Clark

The oil industry has long recognized the need for increasing well productivity. To meet this need, a process is being developed whereby the producing formation permeability is increased by hydraulically fracturing the formation. The "Hydrafrac" process, as it is now being used, consists of two steps: (1) injecting a viscous liquid containing a granular material, such as sand for a propping agent, under high hydraulic pressure to fracture the formation; (2) causing the viscous liquid to change from a high to a low viscosity so that it may be readily displaced from the formation. To date the process has been used in 32 jobs on 23 wells in 7 fields, resulting in a sustained increase in production in 11 wells. INTRODUCTION Need For Process Although explosives, acidizing, and other methods have long been used, there still exists a need for artificial means of improving the productive ability of oil and gas wells, particularly for wells which produce from formations which do not react readily with acids. This paper discusses the development of a hydraulic fracturing process, "Hydrafrac", which shows distinct promise of increasing production rates from wells producing from any type of formation. The method is also considered applicable to gas and water injection wells, wells used for solution mining of salts and, with some modification, to water wells and sulphur wells. Requirements of Process In considering such a possible process, it appeared that certain requirements must be met. Some of these are as follows: A. The hydraulic fluid selected must be sufficiently viscous that it can be injected into the well at pressure high enough to cause fracturing. B. The hydraulic fluid should carry in suspension a propping agent, such as sand, so that once a fracture is formed, it will be prevented from closing off and the fracture created will remain to serve as a flow channel for oil and gas. C. The fluid should be an oily one rather than a water-base fluid, because the latter would be harmful to many formations. D. After the fracture is made, it is essential that the fracturing fluid be thin enough to flow hack out of the well and not stay in place and plug the crack which it has formed. E. Sufficient pump capacity must be available to inject the fluid faster than it will leak away into the porous rock formation. F. In many instances, formation packers must be used to confine the fracture to the desired level, and to obtain the advantages of multiple fracturing. Development of Process As a necessary step in the development of this process, it was deemed advisable to determine if the Hydrafrac fluids were actually fracturing the formation or whether these special fluids were merely leaking away into the surrounding formation. To determine this, a shallow well, 15 feet deep, was drilled into a hard sandstone. Casing was set, the plug drilled, and the well deepened in the conventional manner. A fracturing fluid dyed a bright red was used to break down the formation. Sand mixed with distinctively colored solids was injected into the well with the fracturing fluid to prop open any fracture made in the formation. A simulated gel breaker solution dyed a bright blue was then pumped into the well to determine if the gel breaker would follow the first solution. The results are shown in Figure 1. It was noted that a fracture was formed about the well bore, that the propping agent was transported back into the break, and that the breaker solution did actually follow the fracturing gel out into the fracture. While it is realized that this shallow well test is probably not exactly equivalent to a deep test, the results were interpreted as being a definite indication of what happens down the hole during a Hydrafrac job. Of interest in this connection is an investigation reported by S. T. Yuster and J. C. Calhoun, Jr.' This study, re~orted after the Hydrafrac work was under way, presents some excellent field data supporting the theory of fracturing a formation with hydraulic pressure. METHOD Steps of Hydrafrcu: Process Figure 2 shows a simplified cross-sectional view of a well treated by one version of the process. The first step, formation breakdown, is done with a viscous fluid, usually consisting of an oil such as crude oil or gasoline, to which has been added a bodying agent. Due to availability and price, war-surplus Napalm has been used in the majority of experiments to date. Napalm is the soap which was used in the war to make "jellied gasoline". The next step consists of breaking down the viscosity of the gel by injecting a gel-breaker solution and then after several hours, putting the well back on production. Figure 3 shows diagram-matically, a typical field hookup. The oil or gasoline is unloaded into the 10 bbl. tank shown on the left rear of the truck. This base fluid is picked up by the mixing pump and pumped through the jet mixer, where the granular soap is added. Next it goes into a small mixing tub, from which the high-pressure pump takes suction. The solution is then pumped into the well. The breaker solution is then taken from an extra tank and is displaced into the well immediately following the gel. When required, additional trucks may

Jan 1, 1949

By J. B. Clark

The oil industry has long recognized the need for increasing well productivity. To meet this need, a process is being developed whereby the producing formation permeability is increased by hydraulically fracturing the formation. The "Hydrafrac" process, as it is now being used, consists of two steps: (1) injecting a viscous liquid containing a granular material, such as sand for a propping agent, under high hydraulic pressure to fracture the formation; (2) causing the viscous liquid to change from a high to a low viscosity so that it may be readily displaced from the formation. To date the process has been used in 32 jobs on 23 wells in 7 fields, resulting in a sustained increase in production in 11 wells. INTRODUCTION Need For Process Although explosives, acidizing, and other methods have long been used, there still exists a need for artificial means of improving the productive ability of oil and gas wells, particularly for wells which produce from formations which do not react readily with acids. This paper discusses the development of a hydraulic fracturing process, "Hydrafrac", which shows distinct promise of increasing production rates from wells producing from any type of formation. The method is also considered applicable to gas and water injection wells, wells used for solution mining of salts and, with some modification, to water wells and sulphur wells. Requirements of Process In considering such a possible process, it appeared that certain requirements must be met. Some of these are as follows: A. The hydraulic fluid selected must be sufficiently viscous that it can be injected into the well at pressure high enough to cause fracturing. B. The hydraulic fluid should carry in suspension a propping agent, such as sand, so that once a fracture is formed, it will be prevented from closing off and the fracture created will remain to serve as a flow channel for oil and gas. C. The fluid should be an oily one rather than a water-base fluid, because the latter would be harmful to many formations. D. After the fracture is made, it is essential that the fracturing fluid be thin enough to flow hack out of the well and not stay in place and plug the crack which it has formed. E. Sufficient pump capacity must be available to inject the fluid faster than it will leak away into the porous rock formation. F. In many instances, formation packers must be used to confine the fracture to the desired level, and to obtain the advantages of multiple fracturing. Development of Process As a necessary step in the development of this process, it was deemed advisable to determine if the Hydrafrac fluids were actually fracturing the formation or whether these special fluids were merely leaking away into the surrounding formation. To determine this, a shallow well, 15 feet deep, was drilled into a hard sandstone. Casing was set, the plug drilled, and the well deepened in the conventional manner. A fracturing fluid dyed a bright red was used to break down the formation. Sand mixed with distinctively colored solids was injected into the well with the fracturing fluid to prop open any fracture made in the formation. A simulated gel breaker solution dyed a bright blue was then pumped into the well to determine if the gel breaker would follow the first solution. The results are shown in Figure 1. It was noted that a fracture was formed about the well bore, that the propping agent was transported back into the break, and that the breaker solution did actually follow the fracturing gel out into the fracture. While it is realized that this shallow well test is probably not exactly equivalent to a deep test, the results were interpreted as being a definite indication of what happens down the hole during a Hydrafrac job. Of interest in this connection is an investigation reported by S. T. Yuster and J. C. Calhoun, Jr.' This study, re~orted after the Hydrafrac work was under way, presents some excellent field data supporting the theory of fracturing a formation with hydraulic pressure. METHOD Steps of Hydrafrcu: Process Figure 2 shows a simplified cross-sectional view of a well treated by one version of the process. The first step, formation breakdown, is done with a viscous fluid, usually consisting of an oil such as crude oil or gasoline, to which has been added a bodying agent. Due to availability and price, war-surplus Napalm has been used in the majority of experiments to date. Napalm is the soap which was used in the war to make "jellied gasoline". The next step consists of breaking down the viscosity of the gel by injecting a gel-breaker solution and then after several hours, putting the well back on production. Figure 3 shows diagram-matically, a typical field hookup. The oil or gasoline is unloaded into the 10 bbl. tank shown on the left rear of the truck. This base fluid is picked up by the mixing pump and pumped through the jet mixer, where the granular soap is added. Next it goes into a small mixing tub, from which the high-pressure pump takes suction. The solution is then pumped into the well. The breaker solution is then taken from an extra tank and is displaced into the well immediately following the gel. When required, additional trucks may

Jan 1, 1949

By James W. Dally, William L. Fourney, Anders Ladegaard Peterson

For the past 3 years, the authors have been conducting research sponsored by the National Science Foundation (RANN) to improve the process of excavation by drilling and blasting. The approach followed has been experimental where the development of stress waves and fractures initiated at the bore hole have been investigated in order to obtain a complete understanding of the dynamic fracture process. The second step in the approach has been to introduce modifications in the drill and blast procedure which will permit closer control of the fracture process. The laboratory investigations involve high speed photography where the dynamic fracture process is recorded with a Cranz-Schardin 1, 2 multiple-spark camera. The camera is equipped with 16 spark gaps which are pulsed at 25 K volts to produce an intense but very short (0.5 sec) flash of light. The camera is capable of recording 16 photographs of a dynamic event at framing rates which can be varied from 30,000 to 1,500,000 frames per second. The exposure time is sufficiently short to stop motion associated with detonating explosive charges and to make visible the details of the fracture process at a bore hose. The bore hole in a massive intact rock formation is modelled with a two dimensional plate containing a circular hole to represent the bore hole. The model material employed is a transparent polyester known commercially as Homalite 100.* This polymeric material is extremely brittle as evidenced by its extremely low fracture toughness of [ ]. The fracture toughness is a measure of the ability of a material to resist the propagation of flaws or small cracks. In comparison, Schmidt3 has recently measured the fracture toughness of Salem limestone and determined [ ]. Thus, the Homalite 100 should closely model the brittle nature of rock where fractures occur at small flaws and propagate without any apparent plastic deformation. Homalite 100 is also birefringent, which indicates that it becomes optically anisotropic when subjected to either static or dynamic loads. Circularly polarized light is transmitted through the loaded Homalite 100 model in a polariscope4 and the birefringence produces an optical interference pattern which is called a fringe pattern. For dynamic photoelasticity, the multiple-spark camera is equipped with polaroid filters to produce the circularly polarized light required to generate the photoelastic fringe patterns. An example of a singlespark frame showing a fringe pattern from a typical experiment is presented in [Fig. 1]. The photograph was taken 0.000072 sec (72 sec) after the detonation of the explosive charge. The circular fringes are due to the outgoing dilatational or P type stress wave and travel with a velocity of 85,000 in. per sec (2260 m/sec) in the Homalite 100. The P wave is followed by a second lower velocity stress wave known as the shear or S type wave which propagates at a velocity of 49,000 in. per sec (1245 m/sec). In the local neighborhood of the bore hole, several radial cracks are visible. These cracks propagate at essentially a constant velocity of 15,000 in. per sec (380 m/sec) prior to arrest. The fringes about the crack tips and in the local region of the bore hole are primarily due to the residual gases contained in the bore hole after the explosive charge was detonated. Sixteen frames similar to this one are recorded during the experiment to give full field visualization of the dynamic event at 16 discrete times over its duration. The fringe order number N is related to the difference in the principal stresses of and 02 according to a stress optic law4: [ ] where f0 = material fringe value, and h = model thickness. The wholefield dynamic-fringe patterns provide a basis for simultaneously observing the interaction between propagating cracks and the stresses which drive these cracks. Fracture Control Experiments Improvements in the efficiency of the drill and blast procedures must involve close control of the fracture process following the detonation of an explosive charge in a bore hole. By control it is implied that the number of cracks initiated and the location of each crack on the wall of the bore hole can be specified. Control also, involves orienting each crack and maintaining the crack path and velocity until the specified crack length is achieved. If the entire fracture process can be controlled, then rounds can be designed to optimize volume removed. fragment size and minimize costs. One area of blasting where fracture control is vitally important is in underground excavation where the strength and stability of the rock walls must be maintained and smoothness and precision of the walls must be achieved. The smooth blasting method is one of the most commonly employed procedures for achieving some degree of fracture control. In smooth blasting, the central region of material is first removed, and then the final row of closely spaced undercharged or cushioned holes are fired to remove the final volume and produce a smooth wall. In some instances, unloaded or dummy holes between the loaded holes are recommended to guide the fracture plane. This investigation pertained to an evaluation of 3 features of the smooth blasting process. These included (a) the effect of stress reinforcement on fracture by simultaneously firing 2 charges; (b) the influence of a dummy hole on control of the fracture planes between 2 simultaneously fired charge holes; and (c) the influence of dummy hole spacing on fracture plane control.

Jan 1, 1979

By P. A. Laxen, H. R. Spedden, A. M. Gaudin

WHEN a mineral particle is fractured, bonds between the atoms are broken. The unsatisfied forces that appear at the newly formed surface1 are considered to be responsible for the adsorption of ions at the mineral surface. A knowledge of the mechanism and extent of ion sorption from solution onto a mineral surface is of interest in the development of the theory of flotation.2,3 Study of the adsorption of sodium from an aqueous solution on quartz offers a simple approach to this complicated problem. The availability of a radioisotope as a tracer element meant that accurate data could be obtained.4,5 Three main factors which appeared likely to affect the adsorption of sodium are: 1-concentration of sodium in the solution, 2-concentration of other cations in the solution, and 3-anions present in the solution. Hydrogen and hydroxyl ions are always present in an aqueous solution. By controlling the pH, the concentration of these two ions was kept constant. The variation in the amount of sodium adsorbed with variation in sodium concentration was then determined under conditions standardized in regard to hydrogen ion. The effect of concentration of hydrogen ions and of other cations was also measured. A few experiments were made to get a preliminary idea on the effect of anions. The active isotope of sodium was available as sodium nitrate. Standard sodium nitrate solutions were used throughout these experiments except when the effects of other anions were studied. It was found that sodium adsorption increased with sodium-ion concentration, but less rapidly than in proportion to it. Increasing hydrogen-ion concentration, or conversely decreasing hydroxylion, brings about a comparatively slight decrease in sodium-ion adsorption. Increasing the concentration of cations other than hydrogen or sodium decreases somewhat the adsorption of sodium ion. It would appear as if the kind of anion is a secondary factor in guiding the amount of sodium ion that is adsorbed. Materials and Methods Quartz The quartz was prepared as in previous work in the Robert H. Richards Mineral Engineering Laboratory4 except for the refinement of using de-ionized distilled water for the final washing of the sized quartz, prior to drying5 To minimize the laborious preparation of quartz, experiments were made to determine whether the sodium-covered quartz could be washed free of sodium and re-used. The experiments were successful as indicated by lack of Na' activity on the repurified material and by its characteristic sodium adsorption. Table I gives the spectrographic analyses of the quartz used. The quartz ranged from 16 to 40 microns in size, averaging about 23 microns (microscope measurement), and had a surface of 1850 sq cm per g (lot I), 2210 (lot II) and 2000 (lot III) as determined by the Bloecher method.6 Radioactive Sodium Method of Beta Counting for Adsorbed Sodium: Na22, the radioisotope of sodium, possesses convenient properties.7 It has a half-life of 3 years, thus requiring no allowance for decay during an experiment. On decay it emits a 0.575 mev ß radiation and a 1.30 mev ? radiation. The decay scheme is illustrated in the following equation: [Y Nam S. - 'Net 3 years] The ß radiation is sufficiently strong to penetrate an end-window type of Geiger-Mueller counting tube. This, in turn, makes it possible to use external counting, a great advantage in technique. Furthermore, it permits the assaying of solids arranged in infinite thickness, while assaying evaporated liquors on standardized planchets. The equipment used was standard and similar to that employed by Chang8 The original active material was 1 ml of solution containing 1 millicurie of Na22 as nitrate. This active solution was diluted to 1000 ml. Five milliliters of this diluted active solution was found to give a quartz sample a sufficiently high activity for accurate evaluation of the sodium partition in the adsorption measurements. Also, 1 ml of final solution gave a sufficiently high count for precision on the liquor analyses. The sodium concentration of the diluted active solution was 1.2 mg per liter, so that 6 mg of sodium for 60 ml of test solution and 12 g of quartz was the minimum amount used. The active solution was stored in a Saftepak bottle. Procedure for Adsorption Tests: The method consisted of agitating 12 g of quartz with 60 ml of solution of known sodium concentration for enough time to establish equilibrium between the solution and the quartz surface. The quartz was separated as completely as possible from the solution by filtering and centrifuging. The activity on the quartz and in the equilibrium solution was measured and the partition of the sodium was calculated from the resulting data. The detailed procedure for the adsorption test is set forth in a thesis by Laxen5 In brief, it included the following steps: 1-Ascertainment of linearity between concentration of Na22 and activity measured. 2-Evaluation of factor to translate activity on solid of infinite thickness in terms of activity on an evaporated active film of minute thickness, on the various shelves of the counter shield. 3-Taking precautions to avoid evaporation of water during centrifuging.

Jan 1, 1952

By Kiyoshi Azuma, Hiroshi Kametani

The variation of the dissolution behavior of a ferric oxide with calcining temperature has been investigated. Samples were prepared by thermal decomposition of ferric hydroxide, nitrate, oxalate, and sulfate at low temperature, followed by the calcination in the temperature range between 600" and 1200°C. The samples of eight series and a fine crystalline sample of hematite were dissolved in 1 N hydrochloric acid at 55.2°C and the results are represented on double-log graphs for convenience. It is confirmed that all dissolution courses follouj either the accelerated process or the parabolic process except in the special case of the crystalline hematite which dissolced in accordance with the uniform dissolution of a particle. Examinations of the physical properties of the oxide powders revealed that the surface area measured by the permeability method is strikingly relevant to the dissolution behavior of the oxide. In the previous paper,' detailed data were presented on the effect of the kind of acid, the solution temperature, and the concentration of acid on the dissolution of two ferric oxides. It was also shown that these sam ples dissolved in strikingly different ways. The present investigation was carried out on the dissolution of various calcined samples prepared from various ferri salts by various methods to ascertain the course of dissolution. Pryor and Evans2 pointed out a change of the dissolution rate at around 700°C for a series of calcined ferric oxides prepared from the hydroxide. Several papers374 reported also the dissolution of ferric oxide samples. It seems, however, that a systematic account of the relationship between the dissolution behavior and physical properties of the oxide has not yet been given. This paper presents the variation of the dissolution of the oxide in relation to the calcining temperature and the change of physical properties of the calcines. EXPERIMENTAL Raw materials were prepared by precalcination of ferric hydroxide, thermal decomposition of ferric nitrate, oxalate, and sulfate, and aerial oxidation of ferric chloride vapor, at as low a temperature as possible. The products were crushed, ground, if necessary, and sieved with a 100-mesh Tylor screen prior to calcination, after which the specimens were dissolved in acid solution. The following is a detailed description of the preparation of the samples. Sample H. About 500 g of ferric chloride (guaranteed reagent) were dissolved in 5 liters of deionized water and filtered. Ferric hydroxide was precipitated by addition of the minimum amount of ammonium hydroxide solution, and the precipitate was washed continuously till chloride ion was not detected by silver nitrate solution, and then filtered. The filter cake was dried at 120°C for a week and ground, and the -100 mesh portion was used. Sample S. Ferric sulfate (guaranteed reagent) was pyrolytically decomposed in a crucible at 700°C for 24 hr and the product was sieved. In this case the following calcination was carried out at temperatures over 700°C. Sample B. Commercial ferric oxide (guaranteed reagent). About 15 kg of ferric nitrate were decomposed in a furnace maintained at 800°C for 2 hr. The actual temperature of the decomposition was not measured. The product was crushed and sieved, and the -100 mesh portion was used. Sample N. About 50 g of ferric nitrate (guaranteed reagent) were decomposed in a beaker in a sand bath until a red-brown dense solid was produced. This product was crushed and sieved, and subjected to complete decomposition at 500°C. The precalcined product was again sieved and used. Sample N2.5. Since the decomposition temperature was not controlled for sample AT, a different sample was prepared in a temperature-controlled furnace. The subscript represents the decomposition at 250°C. The product was treated in the same manner as sample N. Sample Nc. Under atmospheric pressure it is prac-tically inevitable that ferric nitrate hydrate melts to form a brown liquid at about 50°C before pyrolysis. For this reason, the salt was first slowly heated under reduced pressure (about 10-3 mm Hg measured in a trap refrigerated by dry ice-alcohol) to achieve dehydration without melting. About 5 hr were required for the dehydration and the partial decomposition. Then the temperature was elevated to 500° C in air for complete decomposition. The relatively porous product was sieved and used. Sample Ov. About 200 g of ferric oxalate hydrate (extra pure) were dehydrated under reduced pressure (as described above) followed by thermal decomposition at 500°C for 6 hr in air. The decomposition of this salt was accompanied by liberation of carbon monoxide, by which the ferric salt was initially reduced to a black powder. The powder changed in turn into brown ferric oxide as the gas liberation decreased and reoxidation predominated. The product consisted of sparkling fine particles passing through a 100-mesh screen. However it was ground and sieved as for the other samples. Sample D. Commercial fine powder for magnetic tape purposes. The preparation was as follows.5 Ferric chloride vapor and preheated excess air were mixed and passed into a reaction tube where oxidation took place at 450°C. The fine powder formed was collected in a cottrell chamber. The product was vacuum-degassed at 450°C for 1 hr and sieved.

Jan 1, 1969

By R. Grierson, L. J. Bonis

The high-temperature Strength oF TD nickel has been observed to be dependent upon the previons thermal and mechanical history of the material. Variations in both the level and the anisotropy of strength have been observed. 01 this paper- these variations are correlated with the storing of annealing resistant elastic strain energy in the matrix of the TD nickel. An x-vay line -broadening tecknique is used to measure the maLrTis elastie strain. THE inclusion of a finely dispersed second phase into a ductile matrix has long been recognized as an extremely effective method of strengthening the matrix both at high and at low homologous temperatures. It has been found, however, that the factors which determine the high-temperature strength are not the same as those which are important at low temperatures. Below 0.5 Tm the size and distribution of the second phase particles are of prime importance in determining the strength,')' while above this temperature the strength is mainly dependent upon the previous thermal and mechanical history of the alloy,3-7 This paper is primarily concerned with explaining the response of the high-temperature mechanical strength of one of these alloys (DuPont's TD nickel) to various thermo-mechanical treatments. It will be shown that this response is not associated with the occurrence of any form of dislocation substructure within the matrix of the alloy. It has been found, however, that a correlation does exist between the elastic strain level in the matrix and the previous thermomechanical history of the alloy and that the observed changes in elastic strain level parallel the measured changes in high-temperature strength. It therefore must be concluded that variations in high-temperature strength are a direct result of the variations in elastic strain level. MATERIAL TD nickel contains approximately 2 vol pct of Tho2 in an unalloyed nickel matrix. It is formed, as a powder, by a chemical technique and this powder is compacted to form ingots which are then extruded to give 21/2-in.-diam rod. Rod of smaller diameter is prepared from the as-extruded rod by swaging. In the studies reported in this paper, 1/2-in.-diam rod was used. This rod received an anneal of 1 hr at 1100°C prior to being used in any of these studies. EXPERIMENTAL TECHNIQUES Two methods were used to examine the structure of the nickel matrix of the TD nickel. These were: 1) transmission electron microscopy; 2) the analysis of the position and profile of X-ray diffraction lines obtained using the nickel matrix as the diffracting media. To prepare thin foils for electron-microscopical examination, slices of TD nickel approximately 0.050 in. thick were cut from the as-received 1/2-in.-diam rod. These were then chemically polished down to 0.045 in., rolled to 0.009 in., given a predetermined heat treatment, and thinned, using a modified Bollman technique, to provide the foils for observation. All observations were carried out at 100 kv, using a Hitachi HU-11 electron microscope. Specimens of the undeformed rod were prepared by grinding down the 0.050-in.-thick slices to approximately 0.015 in. and then thinning chemically and electrolytically to give the thin foils. The X-ray specimens were prepared by rolling 0.375-in.-thick rectangular blocks down to 0.075 in. The surfaces of the rolled material were ground flat, chemically polished to remove the layer disturbed by the grinding, and given a predetermined anneal in an inert atmosphere. They were then ground lightly to check their flatness and given a final chemical polish prior to being examined. The X-ray diffraction line profiles were measured using an automated Picker biplane diffractometer. A special specimen holder was built to allow a more accurate and reproducible positioning of the specimen. The line profiles were determined by carrying out intensity measurements at intervals of either 1/30 deg or 1/60 deg over a range of 3 deg on either side of the nickel peaks of interest. A piece of pure nickel which had been recrystallized to give a large grain size was used as a standard to give the X-ray line profile generated by a strain-free matrix. The analysis of the X-ray diffraction line profiles is a modification of that due initially to Warren and Aver-bach8and has been described elsewhere.3 This analysis gives a measurement of two parameters associated with the structure of the nickel matrix. These parameters are: 1) the size of the coherently diffracting domains within the nickel matrix; 2) the magnitude of the elastic strains in these domains. Both of these parameters are first determined in terms of a Fourier series. These series are obtained from other Fourier series which describe the measured profile of the X-ray diffraction lines. Thus, for both the coherently diffracting domain size and the elastic strain level, it is possible to plot Ft (the Fourier coefficient) against t (the term in the Fourier series), where t can be expressed in terms of a distance L and the Fourier coefficient Ft(S) (associated with elastic strain level) can be expressed in terms of the root mean square strain (e2)1/2. Thus a plot of (F 2)1/2 vs L can be obtained. Plots of this type are shown graphically in Figs. 6 and 8. Interpretation

Jan 1, 1968

By J. Sibakin, R. D. Hindson

In order to devise a practicable soaking pit schedule for use at The Steel Co. of Canada Ltd.'s Hamilton Works, soaking pit heating temperatures, sooking times, pit capacity, and safe maximum mill drafts were correlated with fluctuations in the current or load of the bloom mill driving motor. Other variables such as total delays in the pit, rolling schedules, mill delays, and track times were also investigated. IN order to show an easily applied and accurate means of establishing soaking pit heating temperatures, soaking times, pit capacity, and safe maximum mill drafts, these various factors are correlated herein with fluctuations in the current or load of the bloom mill driving motor. Rolling practices have a considerable influence on the production capacity of a blooming mill. The maximum values of the torque, in particular, are of importance, since even instantaneous current peaks lead to the tripping of the motor by the overload relay and result in loss of mill time. The establishment of safe maximum drafts and accelerations for ingots of different sizes and of a soaking pit practice which would ensure a consistent and satisfactory plasticity of the metal is of considerable importance for increasing the efficiency of both the blooming mill and the soaking pits. The Bloom Mill Dept. of the Hamilton Works, The Steel Co. of Canada Ltd., is equipped with one 44 in. mill driven by a 7000 hp motor with the setting of the overload relay at 22.0 ka. The speed of rotation of the motor is regulated after the Ward-Leonard system. There are three basic speeds of 9.5, 28, and 47 rpm and a further possibility of increasing the speed by weakening the field. This last possibility is hardly ever used during practical operations. The rolling program of the blooming mill is varied, both in the size of the ingots to be handled and in the steel grades. The total tonnage handled by the mill is about 2,000,000 ingot tons per year. At the time of the investigation, the Bloom Mill Dept. was equipped with 22 soaking pits (6 regenerative, 14 bottom-fired, and 2 one-way top-fired pits) with a total bottom area of 2770 sq ft. The pits are fired with a blast furnace-coke oven gas mixture having a calorific value of 155 Btu per cu ft. The foregoing figures show that the production program was such as to impose the necessity of a most efficient usage of the available equipment. For this purpose, the operations of the 44 in. mill and of the soaking pits were investigated, and the results of the investigation were used as a basis for a revised soaking pit schedule and drafting practice. The plasticity of an ingot of a certain chemical composition when being rolled is determined mainly by the following factors: I—the ingot size, both thickness and width; 2—the length of the gas soak; and 3—the surface temperature. The first two factors determine the uniformity of the temperature distribution over the cross-section of an ingot. The third factor introduces the level of the heating of an ingot. The torque produced by an ingot being rolled is determined by the area of the metal displaced, its plasticity, and acceleration values. On the other hand, with shunt motors the torque is determined by the current. This can be assumed to be correct with only a small degree of error for compound motors with a relatively small effect of the series windings as long as the velocity is not regulated by weakening the field. Since the spread is relatively unimportant when compared to the width of an ingot and since it is also reduced several times during rolling by edging passes, the draft alone and not the area of the metal displaced may be taken into consideration with ingots of a similar size. It is therefore possible to determine the main features of the heating and drafting of an ingot by measuring the current and acceleration of the mill motor. After the acceleration has been taken into account, the amount of current will be an indication of how the motor responds to a heating and/or drafting practice and these practices can be adjusted in order to get the desired result. As peak currents are more likely when heavier ingots are rolled, the rolling of plate and slab ingots was investigated. Conditions prevailing when smaller ingots are rolled can be deduced from the results obtained on heavier ingots. All measurements were made when plain carbon grades under 0.15 pct C were rolled. The motor current, the voltage across the armature, and the rpm were recorded simultaneously on synchronized charts, Fig. 1, which moved with the speed of 6 in. per min. Each draft was recorded by a special observer. The rpm curve made it possible to establish the acceleration at any given moment. For purposes of correlation, the maximum current during a pass and the corresponding acceleration were used. The charts made it possible to establish the position of the roller's lever at any given moment as well as the total time of a pass. The slab ingots were divided into three groups (28x35, 28x45, and 27Mx53 in. ingots) and each group was investigated separately. Since they account for most of the current peaks, only flat passes were used for purposes of correlation, a total of 1373 having been investigated.

Jan 1, 1956

By J. E. Campbell, G. D. Calkins, N. M. Ewbank, M. Pobereskin, A. Wesner

GRINDING for size reduction affects the economics of many processes and products. It is essential as the first step in many industrial processes and is also a finishing step for materials with properties depending on particle size, such as talc, cement, and silica sand. Intermediate and fine grinding are vital operations in the U. S. cement industry, which is producing more than 250 million bbl of cement per year.' Wear of the grinding media is a large part of the grinding operation cost. Problems encountered in grinding cement are so complex that evaluation of efficiency and economy of grinding media is difficult.2 It has been especially difficult to evaluate the relative effectiveness of different types of balls because there are no good testing techniques. Many other industrial operations can be evaluated on a laboratory scale with reasonable accuracy. This does not hold true for evaluation of grinding balls. The consistent results obtained in a laboratory test under a given set of conditions are not always borne out in field application. Rough evaluations of the effectiveness of various compositions and types of grinding balls have been made in the field by using a full charge of one type in a mill and comparing the production record with another run using another type of ball. This method is time-consuming and not very precise, as the second run may not have been carried out under identical conditions. Laboratory-scale tests, on the other hand, have yielded inconclusive results, and many investigators have turned their attention to the development of a field testing technique. Field testing small sample lots of grinding balls has been impractical because it is difficult to identify and recover the test specimens from the grinding mill, and individual groups of balls that have undergone different heat treatments can not be separated.".4 To overcome these difficulties, previous investigators have identified the balls by distinctive marks, notches, and drilled holes, but this procedure has three serious drawbacks: 1) Grinding characteristics and quality of the steel balls may be affected. 2) Physical markings may be worn away in the grinding process, especially during a prolonged run. 3) Recovery from the bulk of the charge will be extremely difficult because the markings are hard to see and may be masked by a coating of the product. To circumvent these difficulties, a radioactive-tracer technique was proposed for recovery and separation of steel grinding balls and subsequent evaluation of the various compositions of the balls. The proposed technique involved five basic operations: 1) Thermal-neutron irradiation activation5 of each group of test grinding balls to a different level of specific radioactivity. 2) Addition of groups of radioactive steel-ball specimens into a ball tube mill. 3) Recovery of radioactive steel-ball specimens from the bulk of the mill charge. 4) Separation of the various groups by their specific radioactivity. 5) Evaluation of actual grinding ball wear. Before any physical tests were performed, required neutron irradiation intensity and time were calculated. Probable composition of the steels to be used was ascertained. An examination was made of the radioactive nuclides8 to be formed which would contribute measurably to the radiation level immediately after irradiation and during the test operation. The radioisotopes formed, their types of radiation, and their half lives are listed in Table I. Of these radioisotopes only iron-59 and chromium-51 were significant for the actual wear test. The intensity of radiation that could be detected by a Geiger counter when the test was completed was the basis for the minimum activation level established. The intensity of radiaton that could be safely handled at the beginning of the test was the basis for the maximum activation level, although this was not considered a major problem. Ten groups of grinding balls of various composition and/or surface or heat treatment were to be tested. One group was designated for the minimum irradiation time. The remaining groups were designated for irradiation periods that increased by increments of 33 pct from that of each preceding group. This difference was considered enough for separation and identification of the groups by comparison of specific activity. Potential Hazards: Possible radiation hazards that might be encountered during this experiment were evaluated for the three important phases: 1) the radiation hazard of placing balls and removing them from the mill, 2) contamination of the product cement by radioactive material worn from the balls, and 3) contamination of the steel by the radioactive balls left in the mill. The radiation intensity expected from the whole group of radioactive balls was calculated to be 250 milliroentgen per hr at 1 ft. This meant the balls would require special shielded packaging and warning labels on the shipping containers. In a radiation field of 250 mr per hr a man can work for 1 hr without exceeding maximum permissible weekly exposure. Since the balls could be dumped into the mill in a matter of seconds, relatively little radiation exposure was anticipated at this stage of the operation. If the weight loss in the balls was 7.7 pct per month and the cement feed through the mill was

Jan 1, 1958

By W. N. Poundstone, P. C. Thakur

Introduction Horizontal drilling in coal mines is a relatively new technology. The earliest recorded drilling in the United States was done in 1958 at the Humphrey mine of Consolidation Coal Co. for degasification of coal seams. Spindler and Poundstone experimented with vertical and horizontal holes for several years. They concluded in 1960 that horizontal drilling in advance of underground mining appeared to offer the most promising prospect (for degasification) but effective and extensive application would be dependent upon the ability to drill long holes, possibly 300 to 600 m, with reasonably precise directional control and within practical cost limits (Spindler and Poundstone, 1960). Mining Research Division of Conoco Inc., the parent company of Consolidation Coal Co., began a research program in the early 1970s to achieve the above objective. The technology needed to drill nearly 300 m in advance of working faces was developed by 1975 and experiments on advance degasification with such deep holes began in 1976. Preliminary results of this research have already been published (Thakur and Davis, 1977). To date nearly 4.5 km of horizontal holes have been drilled for advance degasification and earlier results were reconfirmed. In summary, these are: • The greatest impact of these boreholes was felt in the face area where methane concentrations were reduced to nearly 0.3% in course of two to three months from original values of nearly 0.95%. • The methane concentration in the section return reduced to 50% of its original value immediately after the boreholes were completed, indicating a capture ratio of 50%. • The total methane emission in the section (rib and face emission plus the borehole production) did not increase but rather gradually declined with time. • Initial production from 300 m deep boreholes in the Pittsburgh seam varied from 3 m3/min to 6 m3/min but then slowly declined as workings advanced inby of the drill site (well head) exposing a larger surface area parallel to the borehole. Encouraged by these results, it was decided to design a horizontal drilling system that would be mobile and compatible with other face equipment. A mobile horizontal drill can be divided into three subsystems: the drill rig, the drill bit guidance system, and borehole surveying instruments. The drill rig provides the thrust and torque necessary to drill 75- to 100-mm diam holes up to 600 m deep and contains the mud circulation and gas cuttings separation systems. The drill bit guidance system guides the bit up, down, left, or right as desired. Borehole surveying instruments measure the pitch, roll, and azimuth of the borehole assembly. Additionally, it also indicates the thickness of coal between the borehole and the roof or floor of the coal seam. Thus, it becomes a powerful tool for locating the presence of faults, clay veins, sand channels, and the thickness of coal seam in advance of mining. In recent years, many other potential uses of horizontal boreholes have come to light, such as in situ gasification, longwall blasting, improved auger mining, and oil and gas production from shallow deposits. The purpose of this paper is to describe the hardware and procedure for drilling deep horizontal holes. The Drilling Rig [Figures 1 and 2] show the two components of the mobile drilling rig: the drill unit and the auxiliary unit. The equipment (except for the chassis) was designed by Conoco Inc. and fabricated by J. H. Fletcher and Co. of Huntington, WV. The drill unit. It is mounted on a four-wheel drive chassis driven by two Staffa hydraulic motors with chains. The tires are 369 X 457 mm in size and provide a ground clearance of 305 mm. The prime mover is a 30-kw explosion-proof electric motor which is used only for tramming. Once the unit is Crammed to the drill site, electric power is disconnected and hydraulic power from the auxiliary unit is turned on. Four floor jacks are used to level the machine and raise the drill head to the desired level. Two 5-t telescopic hydraulic props, one on each side, anchor the drill unit to the roof. The drill unit houses the feed carriage, the drilling console, 300 m of 3-m-long NQ, drill rods, and the electric cable reel for instruments. The feed carriage is mounted more or less centrally, has a feed of 3.3 m, and can swing laterally by ± 17°. It can also sump forward by 1.2 m. The drill head has a "through" chuck such that drill pipes can be fed from the side or back end. General specifications of the feed carriage are: [ ] The auxiliary unit. The chassis for the auxiliary unit is identical to the drill unit but the prime movers are two 30-kW explosion proof electric motors. It is equipped with a methane detector- activated switch so that power will be cut off at a preset methane concentration in the air. No anchoring props are needed for this unit. The auxiliary unit houses the hydraulic power pack, the water (mud) circulating pump, control boxes for electric motors, a trailing cable spool, and a steel tank which serves for water storage and closed-loop separation of drill cuttings and gas.

Jan 1, 1981

By R. F. Dickerson, D. A. Vaughan, A. F. Gerds

ALTHOUGH the metallurgy of uranium has been under intensive study since the early 1940's, no systematic effort has been made to identify the non-metallic inclusions in uranium. Uranium carbide (UC), which is probably the most common inclusion found in graphite-melted metal, has been tentatively identified by previous investigators, but the other nonmetallic inclusions have received little attention. Since metallography is a valuable tool in metallurgical studies, the metallographic identification of the nonmetallic inclusions in uranium is important. Such an investigation has been completed and the identification of slag-type inclusions and of uranium monocarbide, uranium hydride, uranium dioxide, uranium monoxide, and uranium mononitride is described. Metallographic Preporation It is often possible to prepare specimens for metal-lographic examination equally well by several methods. The specimens which were examined in this work were prepared by one of two acceptable methods. For the convenience of the reader, both methods will be discussed in detail and will be referred to simply as Method I or Method II in the subsequent sections. For both Methods I and 11, specimens for microscopic examination usually were mounted either in bakelite or in Paraplex room temperature mounting plastic. Method I—Specimens were ground in a spray of water on a revolving disk covered successively with 120-, 240-, and 600-grit silicon carbide papers. It was necessary to perform the final grinding operation carefully on worn 600-grit paper to keep the scratches as fine as possible. After washing and drying, the specimens were polished for 3 to 4 min on a slow speed wheel (250 rpm) covered with a medium nap cloth. Diamet Hyprez Blue diamond polishing paste, Grade 00, 0 to 2 µ, was used as abrasive with kerosene as lubricant on the wheel. Specimens were washed thoroughly in alcohol and final polished electrolytically in an electrolyte composed of 1 part stock solution (118 g CrO, dissolved in 100 cm3 H2O) with 4 parts of glacial acetic acid. A stainless steel cathode was used. At an open circuit potential of 40 v dc, a polishing time of 2 sec retained inclusions well with the bath at room temperature. If additional etching was required to sharpen the interface between the metal and the inclusions, an electrolyte composed of 1 part stock solution (100 g CrO3 and 100 cm8 H20) and 18 parts glacial acetic acid was used at room temperature. Best results were obtained by etching for from 10 to 15 sec at 20 v dc in the open circuit. Surfaces obtained by this method are suitable for microscopic examination. However, if desired, they may be etched further with other chemicals. Method 11—Rough grinding was done on a wet 180- or 240-grit continuous grinding belt. The specimen was then ground by hand successively on 240-, 400-, and 600-grit silicon carbide papers in a stream of water. Final polishing was accomplished on a 4 in. high speed wheel (3400 rpm) covered with Forstmann's cloth. Linde B levigated alumina, suspended in a 1 volume pet chromic acid solution, was the abrasive. Specimens usually were polished in 5 min or less by this technique. Often the inclusions present in the metal were identified in the mechanically polished condition. When etching was required to outline inclusions more sharply, one of the two following methods was used. In the first method, the specimen is etched lightly while electropolishing in the chromic-acetic acid solution described above (1 part of stock solution to 4 parts of acetic acid). The electrolyte was refrigerated in a dry ice-ethyl alcohol bath and specimens were etched at 60 v dc on the open circuit for 2 or 3 cycles of 3 to 4 sec each. The second technique utilizes electrolytical etching at about 10 v dc (open circuit) in a 10 pet citric acid solution at room temperature. X-Ray Diffraction Technique The major problem in the identification of inclusions in metals by X-ray diffraction techniques is the extraction of a sufficient amount of each type of inclusion to obtain an X-ray diffraction pattern. In the present study, X-ray diffraction patterns were obtained from individual inclusions of the order of 10 µ diam. The polished and etched samples shown in the micrographs were examined at a magnification of X54 or XI00 with a binocular microscope. This allowed sufficient working distance to extract the inclusions with a needle probe for powder X-ray diffraction analysis. Friable inclusions such as MgF2, CaF2, UO2, and UH3 could be freed from the metal by probing the as-polished and etched surface. The fine particles then were picked up on the end of a Vistanex-coated glass rod (0.002 in. diam) which was held in a brass adapter made to fit the powder X-ray diffraction camera. The end of the glass rod was centered in the path of the X-ray beam. In the case of the UC, UO, and UN inclusions which are smaller in size, more metallic in appearance, and less friable than the other inclusions, it was necessary to etch the inclusion in relief before extraction. UN inclusions etched sufficiently in relief in the electrolytic polishing solution described in Methods I and II by increasing the polishing time. UN inclusions were relief etched by extending the

Jan 1, 1957

By Frederick F. Frick

SPONGE iron as produced at Anaconda is a fine, -35 mesh, impure product, about 50 pct metallic iron, obtained from the reduction of iron calcine at a temperature of 1850°F by use of coke resulting from slack coal. The metallic iron particles are bulky and spongey and precipitate copper readily and rapidly from a copper sulphate solution. Investigation of the treatment of Greater Butte Project, Kelley, ore at Anaconda early showed the desirability of using sponge iron as a precipitant for the copper in solution resulting from desliming of the ore in a dilute sulphuric acid solution. Anaconda had done considerable work on the production of sponge iron in 1914 for use as a precipitant of copper from leach solutions. Some success and considerable experilence were attained at the time. indicating that, sponge iron might be successfully made by a modification of the process used in 1914, a batch process in which an iron calcine was reduced by means of soft coke, resulting from noncoking coal, in a Bruckner-type revolving horizontal cylindrical furnace widely used 50 years ago. The coke and calcide formed the bed in the Bruckner furnace, which was rotated at about 1 rpm. The bed was brought to a temperature of about 1800°F by means of an oil flame over the surface. Although results were reasonably satisfactory, they did not warrant full development of the process at that time. A good deal of work has been done in the last 50 years on the production of sponge iron. The objective in some cases has been the production of a precipitant for copper from solution, but the bulk of the work has been done for the production of open-hearth steel furnace stock. The production of an open-hearth stock presents two problems rather than one: first, producticon of the sponge iron, and second, what is perhaps of equal difficulty and importance, conversion of the sponge iron into a form suitable for use in the open-hearth furnace. So far as is known to the writer, none of the sponge iron processes tried in the past have proved to be economically feasible. However, Anaconda had a combination of conditions appearing to justify an attempt to produce sponge iron which would serve for the leach-precipitation-float process. It was thought that the process used in 1914, if changed to a continuous one, might work out satisfactorily. The following favorable conditions at Anaconda justified the investigation: 1—A sufficient tonnage of good grade iron calcine resulting from the roasting of a pyrite concentrate in one of the acid plants, at substantially no cost. 2—Reasonably cheap natural gas. 3-—The fact that there was no need for production of a high grade product. 4— The fact that there was no need for obtaining a consistently high reduction of' the iron in calcine. A small revolving Bruckner-type furnace about 2 ft ID by 4 ft long was set up for early pilot work at the research building. This pilot furnace showed that a satisfactory product could be obtained at reasonable cost. It also indicated a marked advantage in preceding the reduction furnace with a furnace of similar size and capacity for preheating and roasting out any residual sulphur from the feed. The small furnace was operated for several months, various details of the process were worked out. and sponge iron was produced to supply a pilot LPF plant which treated 300 lb of Kelley ore pel- hr. Later a second pilot furnace 5 ft in diam and 12 ft long inside was set up at our reverberatory furnace building. This furnace confirmed the data of the small furnace and gave a basis for design of the final plant. At Anaconda a pyrite concentrate, running about 48 pct S, is recovered from copper concentrator tailings by flotation. This concentrate is roasted to sulphur of 3 pct or less at the Chamber acid plant. The iron calcine contains about 57 pct Fe and 18 pct insoluble. The iron calcine feed, as mentioned before, is re-roasted and preheated in a reroast furnace preceding the reduction furnace. Both are of the Bruckner type. The reroasted calcine is fed into the reduction furnace at 800" to 1000°F along with 30 pct slack coal. In the feed end of the furnace the volatile is burned from the slack, giving a soft coke which readily serves for reduction of the iron. Hard metallurgical coke will not serve the purpose. since it does not reduce CO readily at a temperature of 1850°F. All indications are that the actual reduction of the iron is accomplished by carbon monoxide below the surface of the bed, which is 30 in. deep at its center. Apparently there is a constant interchange: Fe²O³ + 3CO = 2Fe - 3CO², CO² + C = 2CO Actually iron oxide is reduced by CO at somewhat lower temperature than the 1850 °F used in the process. but this temperature is necessary to obtain a satisfactory rate of furnace production. The furnace atmosphere is generally reducing, and typical blue carbon monoxide flames satisfactorily cover the bed. Gas flames from four 3-in. Denver Fire Clay Inspirator burners are played directly on the bed, which is slowly cascaded by the 1 rpm of the furnace. An excess of coke is necessary to assure maintenance of good reducing conditions in the furnace bed. Part of this coke is recovered for re-use.

Jan 1, 1954

By R. W. Heckel, R. E. Trabocco

The recovery process in 400 Monel filings was followed, principally, by using the Warren-Averbach technique of X-ray peak profile analysis. The deformation fault probability, a, was 0.006 in samples of unannealed filings. a , the twin fault Probability , was approximately 0.002 in samples of unannealed filings. Both a and 0 were found to "anneal out" at 600°F. The effective particle size and mzs strain increased and decreased in the (111) direction, respectively, with increasing annealing temperature. The actual particle size was found to be almost equivalent to the effective particle size. Tile small values of deformation and twin fault probabilities accounted for the similarity in values of the effective and actual particle sizes. Stored strain energy and dislocation density calculations based on rms strain decreased with increasing annealing temperature. The dislocation density decreased from 10" per sq cm in the unannealed filings to 10' per sq cm in the partially re-crystallized filings. The square root of the dislocation density based on strain to that based on particle size indicated a random dislocation distribution in the unannealed filings. The dislocation arrangement changed to one with dislocations in cell walls with increasing annealing temperature. THE recovery processes which occur in metals are generally thought to be a redistribution and/or annihilation of defects.' Investigators' have shown that recovery processes can be characterized by X-ray line-broadening analyses. Michell and Haig4 measured the stored energy of nickel powder by calori-metry and found the value to be greater by a factor of 2.5 than that from X-ray data obtained by the Warren-Averbach technique.= Minor increases in particle size occurred up to 752°F (recovery), while above 752°F the particle size increased greatly due to recrystalliza-tion. X-ray microstrain values decreased between room temperature and 392"F, remained constant from 392" to 752"F, and decreased from 752°F to a negligible value at 1112°F. Faulkner developed an equation for calculating stored strain energy based on X-ray line-broadening data which gave a closer correlation of measured and calculated stored strain energy based on the data of Michell and Haig. The stored strain energy released during recovery is predominately dependent on the decrease in dislocation density which was p-enerated from cold work.7 Stored energy has been measured8 in alkali halides during recovery and recrystallization and 80 pct of the stored energy was found to be released during recovery. Dislocation distributions have been studiedg in a number of fcc metals by thin-film electron microscopy. Howie and Swann" found the stacking fault energy of copper and nickel to be 40 and 150 ergs per sq cm, respectively. ~rown" has pointed out that these stacking fault energy values should be corrected to 92 and 345 ergs per sq cm, respectively. The dislocation distribution of a metal is directly dependent on the stacking fault energy of the system. Metals of high stacking fault energy such as aluminum cross-slip readily and do not form planar arrays of dislocations. Metals of lower stacking fault energy such as stainless steels" do not cross-slip readily. Cold-worked nickel has been found to form a cellular dislocation structure after annealing.13 The relatively high stacking fault energy of nickel and copperlo to a lesser extent favor cellular structures of dislocations rather than planar arrays after deformation. The present study of recovery was carried out on a Ni-Cu alloy (Monel 400) to compare with prior studies for pure nickel and pure copper. X-ray line-broadening techniques were used to measure the effect of recovery temperature on rms strain and particle size and the results were compared with previous studies on copper'4-'7 and nickel., Calculations were also made on stacking fault probabilities, dislocation density, dislocation distribution, and stored strain energy as affected by temperature. EXPERIMENTAL PROCEDURE The nominal analysis of the Monel 400 used in this investigation was: 66.0 pct Ni, 31.5 pct Cu, 0.12 pct C, 0.90 pct Mn, 1.35 pct Fe, 0.005 pct S, 0.15 pct Si. The annealed material was cold-reduced in two batches, one 50 pct and the other 80 pct. It was originally planned to conduct line-broadening studies of these bulk samples; however, rolling textures that developed produced low-intensity peaks which were not suitable for line-broadening analysis. Filings were prepared at room temperature from both the 50 and 80 pct cold-reduced specimens, series A and series B, respectively, and were not screened prior to heat treatment or X-ray studies. Heating to the annealing temperature, 200" to 120O°F, was accomplished in a matter of minutes in a hydrogen atmosphere. Following heat treatment, some of the filings were mounted and polished for microhardness measurements with a Bergsman microhardness tester, using a 10-g load. A G.E. XRD-5 diffractometer using nickel-filtered Cum radiation was used to obtain all diffraction patterns. Only (111)- (222) line-broadenin data were used in the present study since the {400f peaks were too weak to use. The Fourier analysis of the (111) and (222) peak

Jan 1, 1969

By Charles E. Lundin

The character of the Er-D system was established by determining pressure-temperature-composition relationships. A Sieuerts&apos; apparatus was employed to make measurements in the temperature range, 473" to 1223"K, the composition range of erbium to ErD3, and the pressure range of 10~s to 760 Torr. The system is characterized by three homogeneous phase regions: the nzetal-rich, the dideuteride, and the trideuteride phases. These phases and their solubility boundaries were deduced from the family of isotherms of the system zchich relate the pressure-temperature-composition variables. The equilibrium plateau decomposition relationships in the two-phase regions were determined from can&apos;t Hoff plots to be: The differential heats of reaction in these two regions are AH = - 53.0 * 0.2 and -20.0 *0.1 kcal per mole of D2, respecticely. The differential entropies of reaction are AS = - 36.3 * 0.2 and - 31.0 * 0.2 cal per mole D2. deg, respectively. Relative partial molal and intepal thermodynamic quantities were calculated from the pure metal to the dideuteride phase. The study of the Er-D system was undertaken as a logical complement to an earlier study of the Er-H system.&apos; The primary interest was to compare the characteristics of the two systems and relate the difference to the isotopic effect. Studies of rare earth-deuterium systems by other investigators have been very limited in number and scope. Furthermore, there is even less information available wherein an investigator has systematically compared a binary rare earth-hydrogen system with the corresponding rare earth-deuterium system. The available information consists primarily of dissociation pressure measurements in the plateau pressure region of a few rare earths. Warf and Korst&apos; determined dissociation pressure relationships for the La- and Ce-D systems in the plateau region and several isotherms for each system in the dideuteride region. They compared these data with those of the corresponding hydrided systems. The study of these systems as a whole was very cursory and did not give sufficient data for a thorough comparison of the effect of the hydrogen vs the deuterium in the respective rare earths. The heat capacities and related thermodynamic functions of the intermediate phases, YH, and YD2, were determined by Flotow, Osborne, and Otto,~ and the investigation was again repeated for YH3 and YD3 by Flotow, Osborne, Otto, and Abraham.4 This investigation studied only these specific phases. Jones, Southall, and Goodhead5 surveyed the hydrides and deu-terides of a series of rare earths for thermal stability including erbium. They experimentally determined isotherms of selected hydrides and plateau dissociation pressures for deuterides. These data allowed comparison of the enthalpy and entropies of formation of the dihydrides and dideuterides. To date, no one rare earth has been selected to thoroughly establish the complete pressure-temperature-composition (PTC) relationships of binary solute additions of hydrogen and deuterium, respectively. The objective in this investigation was to provide the first comparison of a complete family of isotherms of a rare earth-deuterium system with those of a rare earth-hydrogen system. This would allow one to determine what differences exist, if any, in the various phase boundaries and the thermodynamic relationships in various regions of the systems. I) EXPERIMENTAL PROCEDURE A Sieverts&apos; apparatus was employed to conduct the experimental measurements. Briefly, it consisted of a source of pure deuterium, a precision gas-measuring buret, a heated reaction chamber, a mercury manometer, and two McLeod gages (a CVC, GMl00A and a CVC, GM110). Pure deuterium was obtained by passing deuterium through a heated Pd-Ag thimble. A 100-ml precision gas buret graduated to 0.1-ml divisions was used to measure and admit deuterium to the reaction chamber. The reaction unit consisted of a quartz tube surrounded by a nichrome-wound furnace. The furnace temperature was controlled by a recorder-controller to . An independent measurement of the sample temperature in the quartz tube was made by means of a chromel-alumel thermocouple situated outside, but adjacent to, the quartz tube near the specimen. Pressure in the manometer range was measured to k0.5 Torr and in the McLeod range (10~4 to 10 Torr) to *3 pct. The deuterium compositions in erbium were calculated in terms of deuterium-to-erbium atomic ratio. These compositions were estimated to be *0.01 D/Er ratio. The erbium metal was obtained from the Lunex Co. in the form of sponge. The metal was nuclear grade with a purity of 99.9+ pct. The oxygen content was reported to be 340 ppm and the nitrogen not detectable. Metallographically the structure was almost free of second phase (<i vol pct). A quantity of sponge was arc-melted for use as charge material. The solid material was compared with the sponge in the PTC relationships. They were found to be identical. Therefore, sponge material was used henceforth, so that equilibrium could be attained more rapidly. The specimen size was about 0.2 gr for each loading of the reaction chamber. The procedure employed to obtain the PTC data was to develop experimentally a family of isothermal curves of composition vs pressure. First, a specimen

Jan 1, 1969

By B. H. Bergstrom, Will Mitchell, T. G. Kirkland, C. L. Sollenberger

IN a previous article&apos; the authors outlined a study of the variables in rod milling and also reported data from a series of open circuit grinding tests on a massive limestone in a 30-in. x 4-ft end peripheral discharge rod mill. As a second part of the experimental program, an analysis is now presented for the 30-in. x 4-ft overflow rod mill grinding under identical conditions, except that discharge ports on the periphery of the mill shell have been sealed so that the products from the present series overflowed through a 9-in. diam .opening in the center of the end plate. A variance analysis has been made of the combined data for the two experiments, and performances of the two mills are compared here. Included in the first report&apos; were descriptions of feed preparation, rod mill circuit, instrumentation and controls, and techniques used to evaluate data. Dependent and independent variables were defined, and variance analyses were made to test the relative significance of variables and to establish magnitude of error for the experiment. Significant data were plotted in various combinations, and conclusions were drawn from the graphs. The procedure and analysis in this series of tests follows the first tests and is not repeated. Data from the second series are recorded in Table I. Listed in the first three columns are the independent variables of feed rate (1000, 2000, 3000, 4000, and 5000 1b per hr), mill speed (50, 60, 70, 80, and 90 pct of critical), and pulp density (50, 60, 70, and 80 pct). The dependent variables, Pso, P100, reduction ratio, slope of the log-log sieve analysis curve, power demand, and Bond work index follow. Of these, only the reduction ratio and the Bond work index were analyzed for significance. Production of new surface as calculated from sieve analyses has not been included for this series because of the questionable assumptions that have to be made to satisfy the formulas involved. The large number of products obtained during the runs precluded the use of surface measurement techniques by the gas adsorption methods at this time; however, samples of all products have been stored for future reference. To test the consistency of the reporting of the sieved products, an averaged sieve analysis was calculated from the wet-dry plots obtained from the three product samples of each run. The resulting averaged analysis was plotted and the P80, selected. The relative deviations of the P80&apos;s from each of the three product samples with respect to the P80 of the averaged analysis were then calculated. In only two sets were the relative deviations (6.2 and 9.9 pct) considered excessive. In each of these two sets, one sieve analysis was obviously out of line; hence that analysis was ignored and new averages were computed. This reduced the relative deviations to 1.2 and 2.7 pct respectively. The relative deviations of the product analyses with respect to their averages ranged from 0.1 to 1.4 pct at 1000 lb per hr, 0.0 to 1.1 pct at 2000 lb per hr, 0.2 to 3.0 pct at 3000 lb per hr, 0.3 to 4.3 pct at 4000 lb per hr, and 0.5 to 5.2 pct at 5000 lb per hr. The relative deviation of the 80 pct passing point for 96 dry sieve analyses of the feed with respect to that of the averaged analysis was 7.6 pct. This slightly higher percentage can probably be attributed to a greater proportion of tramp oversize in a crusher product than is ordinarily found in a rod mill product. The last column on Table I lists the adjusted work index, which has been used as the measure of efficiency for the various combinations of operating conditions investigated. Efficiency increases as the index becomes lower. It was reported in the previous paper that the work indexes for the Waukesha limestone used in these experiments decreased as the product size decreased (as calculated from Bond grindabilities). That is, this limestone becomes easier to grind as the material becomes finer. This is unusual, because the work index for most materials as calculated from the Bond grindability has remained constant as the product size decreased or has increased slightly. Table II lists the results of Bond grindability tests at all mesh sizes from 3 to 200 and the work indexes calculated from them. To remove this variation of work index with product size from the data so that results would apply to any material of constant work index, the work index values shown in Table II were plotted against product size on log-log paper. From this curve (a straight line function in this case), the expected work index for the product size for each of the runs of the experiment was obtained. The work indexes as calculated from the reduction ratio and energy consumption were then divided by the corresponding expected work index. The results obtained are reported in percentages on Table I as adjusted work index and are actually percentages of the work index for the Waukesha limestone at the size in question. Multiplication of the work index value for a material of constant index by these percentages should allow the application of the adjusted work index curves to the material. Only the adjusted work index values, not the actual experimental values, were used for the variance analyses and for the graphs.

Jan 1, 1956

By A. A. Collins

When we speak of high copper on a lead blast furnace we think in terms of 4 to 5 pct, or. any lead charge carrying over 1 pct. Any copper on charge will produce its corresponding troubles such as lead well, extra slag losses, drossing problems, and the working up of the dross. This is indeed a very interesting subject and one that used to give the old-time lead metallurgists such as Eiler, Hahn and lles many worries, not so much in the actual operation of the hlast furnace but in the working up of the copper. When the American nletallurgists commenced with the American rectangular-shaped lead blast furnace in the 1870&apos;s and got away from the reverberatories such as were in use in Germany and other parts of the world, they went to greater tonnages, as 80 to 100 tons per day in comparison to the 20 to 30 tons per day in the other processes. With the greater tonnages along with insuficient settling capacity, the silver losses in some cases were increased. Hence the lead-fall was low, for there were no leady concentrates in those days to assist the metallurgist to gain lead or an absorber for the precious metals; and in some cases copper sulphides were added intentionally to the charge to produce a copper matte to lessen the silver losses through the dump slag. The operators in those days thought that where some copper was always present in the lead ores the copper should not enter into the reduced lead and alloy with it. This, by the way, is just the reverse of our present-day practice, when we try to put all of the copper into the blast furnace lead and to remove the same through the drossing kettles. Therefore the furnace was operated to produce a certain amount of matte or artificial sulphides, since, due to the great affinity of copper for sulphur, any copper present would enter the matte almost completely. Thus, the lead bullion produced was practically free from copper. The products of the furnace were metallic lead or lead bullion, containing 05 to 95 pct of the lead and about 96 pct of the silver which were in the ore—a lead-copper-iron matte which contained nearly all the copper in the ore and the slag, the waste product. In the United States, up through the year 1092, we find the small furnace 100 X 32 1/2 in. with 12 tuyeres, some 6 on each side, plagued with a small amount of poorly roasted sulphides— either from heap or hand roasters that produced matte. This matte was roasted and if poor in copper was returned for the ore smelting. Otherwise it was smelted either alone or with additions of rich slags or argentiferous copper ores, the products being lead and a highly cupriferous matte, the latter being subsequently worked up for its copper. The lead metallurgists kept trying and improving on furnace and roasting equipment designs until we find blalvin W. Iles constructing at the old Globe Plant at Denver what came to be the modern furnace. That is, in 1900 he built a furnace of 42 in. width by 140 in. at the tuyeres with a 10 in. bosh and a 16-ft ore column. This type has been more or less standard to the present time, though modified in width and length to meet the demand for large tonnages and improvements in structural details. In 1905 at Cananea, Mexico, Dwight and Lloyd developed the present down-draft sinter machine that has meant so much in producing a well-processed material for the lead blast furnace. In 1912 Guy C. Riddell came forth with double roasting at the East Helena Plant of the American Smelting and Refining Co., which removed the "zinc mush plague." Incidentally, with the introduction of double roasting, which most lead plants were forced into after 1924, when lead flotation came into its own, less matte or no matte was produced. When this stage arrived, the copper was forced into the dross and the casting of lead at the blast furnace lead-wells was stopped. In plants with a fair copper carry 1 pct or better on the blast furnace charge, the lead wells became inoperative once the production of matte stopped. The copper drosses clogged the lead wells and even with bombing, either water or dynamite, the operators could not keep them open. Thus, the lead wells were abandoned in some plants, such as at the El Paso and Chihuahua smelters of the American Smelting and Refinillg Co., and all lead taken out through the first settlers. The elimination of sulphur, espccially sulphide sulphur, from the blast furnace charge and the nonproductiori of matte resulted in a great saving of tinie, energy and equipment in the recirculation of the copper, With the copper content in the dross and dross-fall ranging in quantities from a few percent up to 60 pct, such as at El Paso, a drossing problem was created. As the old-time operators hated dross and buried the same in the shipping bullion, the modern metallurgists from 1925 on decided that with increasing dross-falls they would have to adopt the lead refiner&apos;s ideas of drossing kettles with subsequent treatment of the lead with a sulphur addition to have the shipping lead of 0.01

Jan 1, 1950

By Ottar Dragge, C. Danielsson, Bo Kalling

DISCUSSION, T. L. Joseph presiding L. F. Reinartz (Armco Steel Corp., Middletown, Ohio) —I would like to know, in the practical application of the Kalling process, what kind of a lining was used, how thick was the lining, and how much metal was treated at one time? S. Fornander (author&apos;s reply)—The rotary furnace is lined with a course of fireclay bricks 6 in. thick. This course is backed by 5 in. of insulation. The furnace has a capacity of about 15 tons. Mr. Reinartz—How was the ladle preheated? Mr. Fornander—As pointed out in the paper, the furnace was heated by a gas flame in the beginning of the experiments. During these first tests, however, the desulphurization was inconsistent. We think that this was due to the fact that iron droplets sticking to the furnace walls were oxidized by the gas flame. Now, the furnace is operated without preheating of any kind, and the results are much better. T. L. Joseph (University of Minnesota, Minneapolis, Minn.)—I might add one comment. This furnace was heated with a flame and for a time they had a little difficulty due to some residual metal in the rotating drum that would oxidize in between treatments and they found therefore, that it was very essential to drain the drum completely of metal so that they would not build up any ferrous oxide between treatments and they eliminated some of their erratic heats by maintaining those more reducing conditions. It was interesting to watch this operation. As soon as the drum started to rotate there was considerable flame, at least, at the time I saw it, that came out around the flanges, indicating there was quite a little pressure on the inside of the drum. W. 0. Philbrook (Carnegie Institute of Technology, Pittsburgh)—Is the reaction slag in the Kalling process liquid or solid, and how is it separated from the metal? Mr. Fornander—In the process there is no slag in the usual sense of the word. The lime powder does not melt during the treatment. After the treatment the lime is still in the form of a fine powder. It is separated from the metal by means of a piece of wood of suitable size placed within the furnace before it is emptied. D. C. Hilty (Union Carbide & Carbon Research Laboratories, Niagara Falls, N. Y.)—Dr. Chipman has given us some of his ideas in connection with a specific effect of silicon and silica on sulphur elimination and how silicon might interfere with desulphuriz- ing in the blast furnace. I wonder if he would like to elaborate on the possibility of a similar effect of silicon in the Kalling process? J. Chipman (Massachusetts Institute of Technology, Cambridge, Mass.)—Silicon does not interfere with the Kalling process. Anything that has strong reducing action is good for desulphurization. In these tests where the temperature was low compared to blast furnace temperatures, the silicon that is in the metal is a better reducing agent than the carbon. At high temperatures, carbon is the better. It is not the silicon in the metal that interferes with desulphurization, it is the silica in the slag. Mr. Joseph—I might add that the metal that was tapped from the drum after desulphurization was really at quite a low temperature. It was not measured, but I think it was well under 1300 °C, probably 1200" or a little above that. That was one of the difficulties, and I think there is no question about the fact that the Kalling process—in that it affects desulphurization between powdered lime, solid and liquid iron— is a reaction definitely between the solid lime and the liquid iron. E. Spire (Canadian Liquid Air, Montreal, Canada) — This Kalling process seems very interesting to us and after all it is only a mixing action that is taking place between the iron and the slag. We have attempted to do the same thing in another way. We have placed at the bottom of the ladle a porous plug through which we injected an inert gas. It can be nitrogen or argon. This plug is placed at the bottom of the conventional ladle and gas injected through the plug. That has appeared in our patent. To define this new type of treatment, I use the word gasometallurgy. I do not know if you like it, but it is a way of defining methods of treating metal using gases. What we do is exactly what is done in the exchange process in another way. We have a porous plug at the bottom with a high lime slag on top of the metal. Using this method, we have very good agitation of metal and slag, and with a small flow of gas, we can achieve a very strong agitation. For instance, in the 500 lb ladle, we use only 5 liters of gas a minute. We have an agitation compared to very rapidly boiling water in a pail. Moreover, the agitation can be controlled to create any amount of mixing desired. In a few minutes, with this method, the sulphur dropped from 0.58 to 0.11. These results have been improved since, and we have obtained results like 0.08

Jan 1, 1952

By George M. Schwartz, Ian L. Reid

THE Soudan mine in the Vermilion district of northeastern Minnesota is the oldest iron mine in the state. It has shipped ore every year since 1884 and still contributes a yearly quota of high grade lump ore. No comprehensive report on the Vermilion iron-bearing district has appeared since Clements&apos; monograph,&apos; but Gruner2 discussed the possible origin of the ores in 1926, 1930, and 1932, and recently Reid and Hustad have added data on mining and geology .3, 4 For many years geologists of the Oliver Iron Mining Div., U. S. Steel Corp., have kept up to date a series of plans and vertical sections of the Soudan mine. In connection with mine operation considerable diamond drilling has been done, and this, together with the mine openings, has permitted a reasonably accurate picture of the structure of the orebodies and wall rocks. It has long been evident to geologists familiar with the mine that the ores were not a result of weathering, a point emphasized by Gruner in 1926 and 1930. As the deeper orebodies were developed it also became clear that replacement had played an important part in their development. In recent years it has been recognized that other iron ores were formed by replacement, as Roberts and Bartly5 have argued strongly for the deposits at Steep Rock Lake. On the basis of these facts G. M. Schwartz suggested to members of the Oliver staff that it would be desirable to study the evidence of replacement, particularly the possible alteration of the wall rock which would be expected if the replacement was a result of hypogene solutions. Rock Formations: The formations directly involved in the iron orebodies of the Soudan mine are few though far from simple. The country rock is largely the Ely greenstone of Keewatin age consisting of a mass of metamorphosed lava flows, tuffs, and intrusives which have been more or less altered by hydrothermal solutions. The predominant rock is chlorite schist. Interbedded with the original flows and tuffs are a series of beds and lenses of jasper to which the name Soudan formation has been applied. In the Vermilion district the term jaspilite has been used for interbanded jasper and hematite. According to modern usage these jasper or jaspilite beds do not comprise a formation separate from the Ely greenstone, inasmuch as the beds of jasper are interbedded with the flows and tuffs of the upper part of the greenstone. It would more nearly accord with modern usage to consider the Soudan beds a member of the upper part of the Ely formation. Because of incomplete rock exposure and exploration the number of interbedded jaspilite beds is unknown. In the mine, however, as many as nine major beds of jasper are known on a cross-section of one limb of the syncline, with an equal number on the other limb. In addition diamond drill cores show beds of greenstone down to half an inch in thickness. The thin beds are probably always tuffs. Structure: Rock structure in the Soudan area is complex, and because there are no recognizable horizons within the greenstone it is extremely difficult to work out the details. Generally speaking, the major regional structure is an anticlinorium, the axis trending east-west, with a westerly pitch. The Soudan mine is related to a synclinal structure on the north limb of the anticline about a mile from the west nose of the folded iron formation. The general structure at the mine is that of a closely folded minor syncline on the major regional anticline. A cross fault has dropped the east side so that the bottom of the syncline has not been reached, whereas to the west it is well shown by the mine openings and diamond drill exploration. Throughout the mine the beds of jasper, and ore-bodies that have replaced the jasper, normally dip northward at angles of 80" or steeper. In detail the jasper beds are extremely folded, probably as a result of deformation while they were still relatively unconsolidated. Orebodies: Ore in the Soudan mine is mainly a hard, dense, bluish hematite. Locally ore has been brecciated and cemented by quartz. The vugs commonly occurring near the borders of orebodies are lined with quartz crystals. They seem to have formed as part of the ore-forming process and are evidence that no folding or compression of the ore has taken place. The orebodies are numerous, varying greatly in size. Many lenses of high grade hematite are too small to be mined. Some of the larger orebodies have been followed vertically for as much as 2500 ft and horizontally up to 1500 ft. The large ore-bodies are extremely irregular in outline in the plane of the beds of jaspilite. In width they are more regular, as they are strictly governed by the width of the jaspilite beds and the greenstone wall rock, which seems to have resisted replacement by hematite. At many places the orebodies replace the jaspilite completely and have a footwall and hanging wall of greenstone. At other places either one or both walls may be jaspilite. Geologists who have studied the orebodies in recent years agree that evidence for the replacement origin of the hematite bodies seems conclusive. AS noted above, many of the orebodies replace jaspilite beds from wall to wall with no evidence whatever of compaction. The replacement origin is also supported by details of the banding which is characteristic of the

Jan 1, 1956