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By Paul Gordon, A. S. Iyer

Apreliminary investigation has been carried out on the annealing behavior of explosively deformed copper as compared to that of conventionally deformed copper, using hardness, stored energy, and the sharpening of X-ray spots on Debye rings as the measures of annealing. Some metallographic observations were also made. Two purities of copper were studied. The first was tough pitch copper obtained as a small piece from the guard ring of a sample explosively deformed to approximately 500 kbars.* The second consisted of mentedand deformed these samples quite badly so that only a few small fragments were available for the annealing experiments. As a result, little in the way of quantitatively significant data was gleaned from the measurements. However, several important qualitative differences between the annealing of explosively deformed and conventionally deformed copper have been observed. It is these effects which are described in this report. Fig. 1 presents the results of hardness and stored energy measurements during the annealing of the high-purity copper after explosive and after conventional deformation. The data for the conventionally deformed copper were taken from the work of Gordon;' in this copper it is evident that most of the stored energy is released concommittantly with the hardness drop. As shown in orddon's' investigation, these effect:: are definitely associated with recrystal-lization. In the case of the explosively deformed copper, on the other hand, a large fraction of the stored energy is released before any appreciable hardness change takes place. At 160°C about 40 pct, and at 170°C about. 80 pct, of the stored energy appears to have been released prior to the onset of softening, i.e., undoubtedly prior to recrystallization. In view of the facts that the hardness and energy measurements were made on separate samples and that there is reason to suspect that the intensity of the explosive deformation from sample to sample may not have been uniform, a check run was carried out in which both energq release and hardness changes were measured cnn the same sample. The results of this run, shown in Fig. 2, confirm the conclusion that much of the stored energy is released before the hardness begins to change appreciably. In an effort to obtain some idea of the activation energy associated with the early release of energy, a single callorimetric sample was annealed in two steps, first at 149.5"C and then at 159.5"C, Fig. 3.

Jan 1, 1962

By E. L. Jr. Spain

THE deposits described in this paper occur over much of Wayne County and in the southeast portion of Hardin County, Tennessee, and in the northeast and northwest portions of. Mississippi and Alabama respec-tively. The presence of tripoli in this region has been known for some time and descriptions of various occurrences in Tennessee have been made by Jewell1 and Miser2, while Hilgard3, Lowe4 and Logan5 have treated the deposits that occur in the adjacent portions of Alabama and Mississippi, but there has not been a general survey of the entire field summarizing completely the geology of the deposits with respect to their origin and stratigraphic relationships. During 1934, while engaged in surveys of mineral resources through this region for the Tennessee Valley Authority, the writer collected data with this in mind, and has assembled them in this report. He is indebted to H. S. Rankin and N. A. Rose, of the Tennessee Valley Authority Geology Division, for field assistance.

Jan 1, 1936

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 G. B. Gould

EVERY mathematical statement of a measure of anything (as distin-guished from a count) is followed by a qualification always implied if not explicity stated--that the statement is only an estimate, which may vary from absolute accuracy within certain limits, according to the accu-racy of the measuring instrument. This basic fact, so obvious to engineers, is not so well understood by many that make use of laboratory coal tests. The tolerances that must be allowed in measurements of coal quality are very important in both their commercial and engineering applica-tions-and the significance of these tolerances depends upon the use made of the laboratory coal test. Only in recent years has there been any intensive study of this ques-tion of tolerances as applied to measurements of coal quality. There is very little in our published literature on the subject, and what has been published is not, readily comprehended by those who sell coal, or buy coal and use it, unless they happen to have made a study of the subject them-selves. There is great need for more widespread understanding of what has been discovered and how it applies to the use of laboratory coal tests. What follows is an attempt to condense into a few pages the results of the major investigations so far reported, and to explain their meaning in simple nontechnical language, with the addition of some recent illustrative experiments.

Jan 1, 1937

By E. W. Murbach

The oxidation of uranium carbide by oxygen at various pressures, and by air, has been investigated at temperatures up to 600°C. Arc-melted and cast uranium carbide displays oxidation behavior that appears to be dependent on its treatment after melting. ccReactive" uranium carbide, which results from exposure of freshly melted material to laboratory air, ignites in oxygen at about 275° to 350°C depending on pressure, and in air at about 350°C. The product of the initial oxidation appears to be UO2+.. Continued heating produces U30,. The initial oxidation rate is poportional, roughly, to the square root of the oxygen pressure. The activation energy determined for the initial oxidation reaction is 7.1 kcal, from 400° to 600°C in 10 mm of oxygen. AN investigation of low-decontamination methods for reprocessing uranium carbide reactor fuels is being carried out at Atomics International. One of the reprocessing cycles includes oxidation of the fuel at relatively low temperature (300° to 500°C) and reaction of the product oxide with carbon at elevated temperatures (above 1500°C) to produce uranium carbide. A number of small-scale experiments have been carried out in which uranium carbide was oxidized in air and oxygen at various pressures to measure reaction rates and to obtain other information concerning the oxidation process. In initial experiments with arc-melted UC, ignition was observed at about 275°C in oxygen or 350°C in air. However, similar experiments by Neymark' at 426°C in air showed no ignition. It was soon determined that the difference in oxidation behavior was due to the treatment of the UC following arc-melting. Neymark's experiments were carried out with freshly melted UC while those reported in the

Jan 1, 1963

By W. C. Elmore

THE magnetic powder method, long used for roughly mapping mag-netic fields, has recently been refined 1,2 for investigating the microscopic variations in the surface magnetization of ferromagnetic crystals. Bitter2 first discovered regular powder patterns connected with the crystal structure of such specimens. Several other investigators3 have used Bitter's original technique, with slight modifications, to study the varia-tion of the patterns with variations in the magnetizing field applied parallel to the surface, and their dependence upon the state of strain, of polish and of crystal cut. All of the explanations-of these patterns on metallic ferromagnetics have been couched in general terms. In a preliminary report we have described an improvement in the powder technique which led us to the discovery of a new effect throwing con-siderable light on the magnetic secondary structure of alpha iron4 and of nickel5 single crystals. We now report additional evidence confirming the existence and stability of a magnetic block secondary structure.

Jan 1, 1935

By W. C. Elmore

THE magnetic powder method, long used for roughly mapping mag-netic fields, has recently been refined 1,2 for investigating the microscopic variations in the surface magnetization of ferromagnetic crystals. Bitter2 first discovered regular powder patterns connected with the crystal structure of such specimens. Several other investigators3 have used Bitter's original technique, with slight modifications, to study the varia-tion of the patterns with variations in the magnetizing field applied parallel to the surface, and their dependence upon the state of strain, of polish and of crystal cut. All of the explanations of these patterns on metallic ferromagnetics have been couched in general terms. In a preliminary report we have described an improvement in the powder technique which led us to the discovery of a new effect throwing con-siderable light on the magnetic secondary structure of alpha iron4 and of nickel5 single crystals. We now report additional evidence confirming the existence and stability of a magnetic block secondary structure.

Jan 1, 1936

PRODUCTION of nonmetallic mineral raw materials used by the fertilizer, ceramic, chemical, and construction industries and many others, was greater during 1953 than in any previous year. Production of fertilizer minerals was the highest in history despite somewhat decreased consumption in the second half of the year. Marketable phosphate rock production increased 10 pct to an estimated total of 13.1 million LT, valued at $78.5 million, and potash production was estimated at 1.8 million tons of K2O equivalent, valued at over $60 million. Crude gypsum output exceeded the 1952 total and approached the record of 8.66 million tons mined in 1951, with indications that 1953 will show a new all-time high. Gains and losses among gypsum products were irregular, with major gains in plate- glass and terra-cotta plasters, sanded building plasters and miscellaneous building plasters, and principal declines in base-coat plasters, prepared finishes and tile. Lath and wallboard outputs were each about 20 pct higher than in 1952.

Jan 2, 1954

By R. Crnojevich, J. L. Blanco, C. Nikolic, W. G. Sherwood

This paper describes the process used for cobalt purification and recovery at the AMAX Nickel Refining Co. plant In Port Nickel, Louisiana In the first stage, cobaltic precipitate produced in the Outokumpu electrolytic nickel cobalt separation process is dissolved using methanol and sulfuric acid liquor produced in this step is converted to an ammoniacal complex, and sufficient ammonium sulfate is added to facilitate cobalt oxidation to cobaltic pentammine Nickel is then selectively precipitated as nickel-ammonium sulfate double salt at pH 2 5 Final purification of the pentammine liquor is effected by pH adjustment with ammonia and removal of divalent nickel by ion exchange Purified liquor is then advanced to hydrogen reduction to produce cobalt metal Although similar to other cobalt refining processes used in the nickel industry, the Port Nickel process is unique because of a single point chemistry adjustment for a number of unit operations that occur in succession.

Jan 1, 1981

By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

By Michael E. Zellars

The environment for deposition of marine phosphorite is controlled by the relationship of prevailing climatic, oceanographic, structural, and sedimentary conditions. The set, or sets, of conditions that has evoked the accumulation of economic concentrations of phosphorite has been the subject of a number of previous works. Phosphorite presently considered ore is generally in excess of 10% P205 with an extractable phosphate product in excess of 30% P2O5. The concentration of the product in the ore, its amenability to state-of-the-art extraction, and thickness of the overburden are major controlling factors in determining its mineability. A recent survey of phosphorite deposits in the south-eastern United States included a review of the depositional and structural setting for the major areas of occurrence. The application of previous work regarding phosphorogenesis has led to the establishment of a general depositional model for most mid-Miocene primary phosphorites, and, is also possibly applicable to the Bone Valley (Pliocene-Miocene) district of central Florida.

Jan 12, 1978

By D. Turnbull, J. H. Hollomon, J. C. Fisher

Application of the concepts of nu-cleation and growth to the analysis of experimental transformation data has led to valuable descriptions of phase transformations, an outstanding example being the transformation austenite —* pearlite which has been examined with particular care by Mehl and co-workers.&apos;-5 In addition to the pearlite transformation, the proeutectoid fer-rite and proeutectoid carbide transformations are known to proceed by nucleation and growth. Martensite, on the contrary, until recently was thought to form by a mechanism involving neither nucleation nor growth; however, extension of standard nucleation theory6 suggests that martensite, bain-ite, and the other products of austenite decomposition all grow from nuclei in the parent phase. The theory that martensite forms by nucleation and growth is strongly supported by recent experimental work of Kurdjumov and Maksimova.7 The concepts of nucleatioli and growth have been fruitful also in providing a sound basis for quantitative theoretical treatments of the kinetics of phase transformations. For example, Volmer and Weber8 and Becker and Döring9 developed the theory of nucleation from fundamental considerations to a point where excellent agreement was obtained with the results of experiments on the condensation of supercooled vapors. As a result of their analysis, the kinetics of vapor-liquid transformations now can be predicted. It seems probable that application of the theories of nucleation and growth to a quantitative study of austenite decomposition similarly will clarify the nature of the individual transfor: mations involved, and will enable the calculation of austenite transformation kinetics. In the present paper, the theories of nucleation and growth are applied to the austenite ? martensite transformation in steels. The analysis begins with a discussion of nucleation in single component systems. Martensite appears to be coherent with the parent austenite, hence the nucleation theory is modified to include the effects of elastic distortion. Nucleation in the two component iron-carbon system then is discussed, for most steels are primarily alloys of these two elements. Finally, M. temperatures and martensite transformation curves are calculated for each of several alloy steels of varying carbon and chromium content, and are compared with those determined experimentally by Lyman and Troiano10 and Harris and Cohen.11 Nucleation Theory NUCLEATION IN SINGLE COMPONENT SYSTEMS6,12-14 The work required for reversible formation of a region of phase within the parent a phase is expressed conveniently as the sum of two terms: W1 = Aa, the product of the area of the interface and the interfacial free energy, and W2 = VAf, the product of the volume of the region and the free energy increase per unit volume associated with the transformation. The total work is therefore W = Aa + VAf. When a is more stable than ß, Af is positive and W increases without limit as the volume increases. The transformation a ?ß does not occur. It is nevertheless true that small regions of phase ß enjoy temporary existence here and there in the a. The equilibrium number of ß regions of given size is proportional to exp(— W/kT) per unit volume of a, assuring that larger (ß regions occur with diminishing probability. When a is less stable than ß, Af is negative and W passes through a maximum as V increases. Assuming for simplicity that regions of ß are spherical, as is true when the interfacial tension is isotropic and there are no elastic strains, W = 4r2a + (4/3)*r3Af The maximum value of W is W* = 16iro3/3Af2 [1] for regions with radius r* = -2o/Af. [2] For single component condensed systems it has been shown14 that the steady rate of nucleation of 0 per unit volume of untransformed a is nearly proportional to exp[- (W* + Q)/kT] where Q is the activation energy for the unit processes of adding or removing one atom from an embryo or nucleus. If To is the temperature at which a and ß are in equilibrium, the rate of nucleation is a maximum at a temperature 0 < T < To where (W* + Q)/kT is a minimum. P regions smaller than critical size are called embryos; they tend to grow smaller and disappear, only exceptionally growing larger. Regions equal to or larger than critical size are called nuclei. A critical size nucleus may grow indefinitely large or may shrink back to a, either process decreasing the free energy of the region.

Jan 1, 1950

By C. A. Stewart

PAGE. I. Introduction............ 241 II. Location. TopoGRaphy. and History....... 243 III. DescRiption of the Rocks......... 245 1. Altered Sediments.......... 245 2. Alaskite........... 246 3. Alaskite.Porphyry.....255 4. Quartz Dikes........... 257 5. Biotite.Granite.......... 256 6. Quartz.Porphyryl......... 260 7. Andesite........... 262 8. Miscellaneous Dikes......... 263 IV. Age oF the Rocks and Comparison with other Localities.. 263 V. AlteRation............ 265 1. Sericitization........... 265 2. Silicification........... 266 VI. StRuctuRes........267 1. Dip of the Limestone......... 267 2. Faulting........... 267 VII. Contact - metamorhism,.......... 269 1. Endomorphism.......... 269 2. Exomorphism......... 271 Recrystallization and Loss of Color...... 271 New ,Minerals.......... 272 Addition of Material........ 274 Volume.Changes.... 276 Locus of Metamorhism........ 277 Method of Precipitation.........278 VIII. ORe-Deposits............ 278 1. Contact-Metamorphic Copper.Ores,..... 279 General..........279 Shape and Distribution........ 279 Minerals........... 281 Secondary Enrichment........ 282 2. Disseminated Copper.Ores........ 283 3. Origin of the Copper.Orcs........ 284 4. Lead-Silver Ores.......... 287 IX. Bibliography............ 288

Jan 1, 1913

By Victor deWolfe

INTRODUCTION The Henderson Mine, located near Empire, Colorado, utilizes a continuous panel caving system to extract ore as one of the world&apos;s major producers of molybdenum. Any mine using a caving-by-gravity technique of mining must rely on closely controlled draw of the caved ore. This control is essential to insure proper caving action, to avoid damaging load concentrations of weight and to minimize the dilution of ore with waste material. Henderson is no exception. Draw control is a major factor in all production planning, from long- range plans to short-range and day-to-day ore scheduling. Draw control is reviewed constantly and administered daily in an effort to optimize production efficiency, ore recovery, and cave management. MINING METHOD The cave at Henderson is massive, moving slowly through large panels that are 244 m (800 ft.) wide by 610 m (2,000 ft.) long. Generally two cave areas are drawn at one time. The areas under active draw vary in size but can be as large as 244 m (800 ft.) by 244 m (800 ft. ) containing 400 draw points. Each draw point contains 45,360 mt (50,000 st) on the average and takes about two and one half years to exhaust. A complete panel is worked for seven to ten years. No pillar exists between panels, but rather a buffer zone of broken ore, or "static face," is left in each panel to be drawn with the adjacent, yet-to-be-caved panel in efforts of minimizing dilution of a working area from an exhausted one. (Figure 1) Production drifts are driven on 24.4 m (80 ft.) centers through the ore body. Between the production drifts are funnel-shaped draw bells on 12.2 m (40 ft.) x 24.4 m (80 ft.) centers to receive ore from the cave. Each bell is accessed by two draw points, one from the production drift on either side, thus forming a 12.2 m (40 ft.) x 12.2 m (40 ft.) draw pattern. Extraction of the ore is by rubber-tired, 3.8 m3 (5 yd3) load-haul-dump equipment. The LHDs then tram the ore a maximum of 49 m (160 ft.) to ore passes. Cave initiation and bell development are done from the undercut drifts which are parallel to and 17 m (55 ft.) directly above the production drifts. Longhole rings are drilled and blasted from the undercut drifts to define the bells and establish the undercut for caving. (Figure 2) DRAW CONTROL Since the cave line at Henderson is constantly advancing, it is necessary to be continually initiating new cave at one end while exhausting it at the opposite end. There must exist, therefore, an angle on the ore-waste contact in the broken rock from initiation to exhaustion. The basic concept of draw control is to keep this angle as smooth and even as possible, particularly at the time of exhaustion. If this is achieved, draw points are exhausted more or less in a line, avoiding pockets of remaining ore surrounded by exhausted areas. These pockets would cause spotty ore extraction at the time of exhaustion, increasing the amount of dilution occurring while introducing the potential for significant weight problems in the production area. To arrive at the desired angle on the ore- waste contact, maximum tonnage percentages are assigned to each row of draw points increasing at 10% or 15% increments (depending on cave size and velocity of draw) working away from the cave line. The available tonnage indicated by these percentages is the maximum allowable tonnage to be extracted from each draw point until the available tonnage percent- age is increased. As the cave moves, these percentages increase for each draw point regularly. However, in general the tonnage drawn from each draw point is kept at about 50% of this allowable maximum in order to maintain adequate available tonnage in the cave to sustain production for seven months if cave initiation were to cease. This available tonnage cushion is a safeguard built into the draw control program at Henderson to accommodate fluctuations in the rate of cave advance. When draw points move past the row of 100% tonnage availability, they are drawn past the desired 50% at the same increments per row until exhausted. (Figure 3) To achieve proper draw control, the number of LHD buckets to be taken from each draw point is assigned daily. The actual buckets taken, which may at times deviate from the

Jan 1, 1981

By T. M. Olson, B. G. Lewis

The adsorption of selenite on bentonite, kaolinite, and an iron--coated illite was studied to determine the importance of several soil variables on selenite removal from uranium mill tailings pond seepage. Adsorption capacities of the clays of selenite varied by orders of magnitude and decreased in the following order: illite > kaolinite > bentonite. Maximum adsorption of selenite on the clays was observed between pH 2.5 and the clay&apos;s zero-point-of-charge. Substantially less selenite was adsorbed on these clays in actual mill tailings solutions than in model systems. Virtually all selenite was removed from the tailings at pH 7. The principal removal mechanism at this pH is suggested to be the adsorption of selenite on freshly precipitated metal hydroxides. It is concluded that selenite can be removed from tailings seepage in soil or in clay liners in the presence of kaolinitic clays and iron oxides. Selenite would also tend to be removed from tailings seepage in calcareous strata.

Jan 1, 1984

By Thomas P. Forbath

THE sudden realization that known sulphur reserves amenable to mining by the Frasch hot water process are nearing exhaustion focused attention on widely scattered surface deposits throughout the world. These deposits are not necessarily of lower sulphur content than ores located underneath Louisiana or Texas salt domes which usually aver- age about 30 pct sulphur disseminated in limestone matrix. Their near surface occurrence, however, renders exploitation by the Frasch process impossible. As is well known, the Frasch process depends on the presence of 500 to 1000 ft of overburden and cap rock above the sulphur deposits to permit melting underground sulphur in place by diffusing hot water under pressures of 200 to 600 psig in the formation and raising the molten sulphur to surface by air lift. This process renders possible the production of pure sulphur which is 99.5 pct pure without any subsequent treatment.

Jan 9, 1953

By Leslie Burris, J. C. Hesson

The diffusivity of uranium in liquid cadmium has been measured as a function of temperature by the capillary -bath technique. The diffusivity measurements were made at 4509 500°, 575º, and 650°C, over which temperature range the diffusivity of uranium varied from 1.3 to 2.3 x 10-5 sq cm per sec. The diffusivity of uranium in cadmium was less than that predicted by the Einstein-Stokes equation. VERY little information is available on the diffusivity of uranium in liquid metals. Bonillal has reported a single measurement of the diffusivity of uranium in bismuth. Measurement of the diffusivity of uranium in cadmium was undertaken, not only to supply data in an area in which a dearth of information exists, but also because of a need for this physical property measurement in connection with mass transfer studies now underway in the U-Cd system at Argonne National Laboratory. The capillary-bath method used in this work has been successfully used by other workers2"4 to obtain diffusivities in liquid metal systems. Although considerable experimental technique is required to obtain reliable diffusivity measurements, the basic information required is only that of temperature, capillary tube length, the initial and final compositions of the alloy in the capillary tube, and the composition of the bath into which the solute element is diffusing. From these measurements, diffusivities may be calculated using Fick&apos;s law. The diffusivity equations are presented in the Appendix. Equations which are often used to predict diffusivities in dilute solutions5 are the Einstein-Stokes equation: D= (kT)/(6pµr) and the Eyring equation: D = (?,kT/?y?zp), where D= diffusivity (sq cm per sec), k = Boltzman&apos;s constant, T= temperature ("K), p = viscosity (poises), r = radius of diffusing molecule or species (cm), and A,, Ay, and A, are the diameters of the diffusing molecule in the three axial directions. For round molecules, the Eyring equation gives values which are 37r times larger than the Einstein-Stokes equation. In this work, the experimental values were compared to the Einstein-Stokes equation. Experimental values of diffusivities in liquid metal systems usually lie between those predicted by the Einstein-Stokes and Eyring equations. The viscosity values for cadmium were obtained from the Liquid Metals Handbook.6 EXPERIMENTAL APPARATUS AND PROCEDURE The general procedure employed consisted of rapidly immersing in a molten cadmium bath containing a negligible amount of uranium a capillary

Jan 1, 1963

By John J. Stokes

Literature on the electrolysis of aluminum from cryolite melts and on the structure of these .melts is surveyed critically. Data on density, freezing point, and other properties are reviewed. Theories of the electrolysis are examined in the light of these data. Two theories are presented which account equally well for the observations. In the early days, the process was thought of as a simple one. The actual mechanisms must be tremendously more complex. This paper is devoted to a rather narrow part of the whole subject, namely, what ions are present in the bath, and by what mechanism the current is carried. Using cryoscopic data, Rolin has determined the following composition for molten cryolite. Na2,AlF 3Na + 3F + AlF [1] He postulates the following equation for the ioniza-tion of alumina at low concentrations Al2O3 Al + AlO2 +O. [2] At concentrations above 0.025M he supposes the ionization of alumina to proceed as follows Al2O3 AlO + AlO2. [3] Rolin&apos;s mechanism of electrolysis for the more concentrated solution in Eq. 3 is as follows. The AlO&apos; ion migrates to the cathode. where it is dis- charged, giving metallic aluminum and 0&apos; ions which migrate to the anode. The AlO2- ion migrates to the anode, where it is discharged, giving Al ions and oxygen which reacts with carbon to form CO2 In the case of the extremely dilute solution in Eq. 2 the Al ion migrates to the cathode and the AlO2-and 0 migrate to the anode as before. Another possible mechanism for the electrolysis follows. Ionization of Cryolite Na3AlF4 = 2Na + 2F + NaAlF, Na + AlF-AlF+Na + F [4] Addition of Alumina 2Na2AlF+ 2AL2O4 6Na + 3A12O3F4 [5a] Al2O2F 2AlOF2 [5b] Cathode Process AlF, + 3e+ 4F + Al. [6] Anode Process AlOF2 + 2F AlF1 + 1/20, + 2e or A12O2F + 4F-2A1F; + 0, + 4e. [7] Although both Rolin&apos;s and the present mechanism of the Hall process may be qualitatively satisfactory, present data are not adequate for quantitative tests of the theories.

Jan 1, 1959

By T. Cuscino, D. D. Lane, C. Cowherd, T. E. Baxter

Three types of commercially-available dust suppressants, commonly used in the surface mining industry to control fugitive particulate emissions from unpaved roads, were tested in the laboratory for cohesive strength and resistance to reduced effectiveness by large water addi¬tions. Results indicate that each product tested varies in cohesive strength with a range of 34-300 kPa (5-44 psi). Total cohesive strength depends on the aggregate moisture content and the means by which the sample is dried. Results show that laboratory testing can be used as a decision¬making tool concerning the type of dust suppressant for a particular mining operation. Also, the road surface aggre¬gate moisture content u critical to successful dust suppres¬sant performance. Large water additions to the aggregate road surface after initial suppressant application can pro¬duce nondesirable results. A modified wet sieving analysis is presented to test the effect of water addition.

Jan 1, 1984