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By E. F. Thomas

THE late George S. Rice was active in the inves--I- tigation of bumps, particularly in the last ten years of his career as chief mining engineer of the U. S. Bureau of Mines. Since most of his investigation was carried out in Great Britain, continental Europe, and—to a lesser extent—Canada, his thinking on prevention was influenced considerably by the experience of those countries. It is not surprising, therefore, that when he was called upon a few years before his retirement to investigate bumps in the U. S. and suggest ways to prevent them, he turned to longwall mining. A longwall method had been most successful in combating the bump hazard in mining coal under deep cover, especially in Great Britain, but the prevailing method there at the time was advancing longwall mining, which he knew was uneconomical under U. S. mining conditions. For this reason he proposed a modified retreating longwall system that he believed included the best features of the advancing method. As brought out by Rice,' if the cover is 2000 ft and 50 pct of the coal is extracted, the static load on the remaining pillars will be about 4000 psi, which exceeds the ultimate crushing strength in most instances. If the pillar coal is overloaded before a pillar line is established, then the abutment zone preceding a line of extraction is no place to split pillars or extract them by any method other than an open-end system. Rice therefore advocated open-end mining, preferably by longwall, but he was willing to compromise with long-face mining if the longwall method was not acceptable. Rice's system was put into operation in a mine in Harlan County, Kentucky,3 but subsequent experience has shown that it did not take into account two important factors—avoidance of pillar-line points and maintenance of adequate development in advance of the pillar-line abutment area. For ten years after Rice's retirement the USBM did little investigation and research on bumps, chiefly because so few were occurring that there was not much cause for alarm. But in 1951 there were three occurrences involving fatal injuries, and the Bureau began a statistical survey in that year. C. T. Holland, head of the department of mines at Virginia Polytechnic Institute, was retained as a consultant. The resulting study' of 117 case histories brought out these important conclusions: 1) Almost invariably the bump occurred in a locality affected by the abutment zones of one or more pillar lines. 2) In most cases the locality of the bump was influenced by the abutment zones of more than one pillar line. The term pillar-line point has been used for many years in the Appalachian region for such a situation. Point is used in the geographical rather than the mathematical sense. 3) In pillar-line extraction the following practices are safest in preventing bumps: a. The mine layout should provide for pillars of uniform size and shape along the extraction line. b. The mine layout should be planned so that no development need be done in the abutment zone of a pillar line. c. The layout should permit open-end extraction of pillar lines from the next goaf, so that it will not be necessary to resort to pocket mining, splitting pillars, or any practice that will involve driving in the direction of the goaf within the abutment zone. d. Pillars should be large enough to support area without undue roof and floor convergence before establishment of a pillar line. These are, of course, generalities, and while they are useful in laying out areas where bumps can be expected, they are of limited help in many mines that were committed to a system of mining before it was realized that they were subject to bumps. Under such conditions it becomes necessary to choose between the following alternatives: 1) Abandon the territory, except for pillars that offer no extraction problems. 2) Through experience select the pillars that are most heavily loaded, and, by augering, induce bumps from a safe vantage point so that impinged loads are relieved. This method was first developed at the Gary, W. Va. mines of U. S. Steel Corp. and later adapted to mining thick coal beds at Kaiser Steel's Sunnyside mine in Utah. No scientific method is available to determine where to drill within a loaded pillar. Although this method of unloading has worked very successfully at Gary—with one exception—

Jan 1, 1959

By Fred Leighton, Wilson Blake

The microseismic method of detecting instability and high-stress zones in underground mines was developed by the U.S. Bureau of Mines (USBM) in the early 1940's.l,2 For about 25 years this method has been used throughout the world by mining and construction companies in assessing the stability of underground and surface excavations. Rock noises are the release of small amounts of stored strain energy propagating outward from their source as a seismic wave that can be detected with special geophysical equipment. This release of strain energy is caused by a rock mass undergoing load changes resulting in small-scale displacement adjustments such as microcracking, shearing, sliding, and crushing of crystal grains, some even as small as the elastic displacements themselves. No detailed field studies of any of these mechanisms have been made, and knowledge about the origin of rock noises is relatively incomplete. Laboratory tests have shown that when rocks are stressed in a testing machine, the rock noise rate-the number of rock noises generated per unit of time increases with increasing applied stress.3 This increase is most pronounced as the ultimate strength of the rock is approached. The standard field procedure has been to monitor a rock structure at intervals and plot rock noise rates vs. time, thus gaining a semiquantitative estimate of the behavior and stability of the structure. The success of such monitoring, however, has depended almost entirely on the experience and knowledge of the operator, his interpretation of the data, and his equipment. Very little study has been devoted to the basic properties of the vibration record of a rock noise. USBM decided that the microseismic method was being used sufficiently by various mining and construction companies

Jan 1, 1970

By Harry V. Warren

During the past few years the author has had an opportunity to examine a number of base-metal mines in the western United States, France and Spain. Nearly all of these mines produced some silver, and laboratory studies of the relationship between the silver and the various base-metal minerals in the ore have given some interesting and, it is to be hoped, enlightening facts. As this work was undertaken primarily to ascertain the amount of silver chemically associated with, not controlled by, the various base-metal minerals, the laborious and not always satisfactory method of picking out "pure" samples of each mineral from many localities was employed. However, the results obtained have justified the time that has been spent. Not only have the conclusions arrived at in different mining areas been in accordance with one another but in many instances they have afforded scientific support or explanations for much of the knowledge that operators in the various camps have acquired by practical experience. Although many of the conclusions are not new, many of the facts are, and the results, if they do nothing else, serve to confirm the findings of some of the other workers in this field, particularly the work of Nissen and Hoyt1 and of F. N. Guild2. Dr. S. G. Lasky3 has suggested a novel method for determining from a series of algebraic equations the amount of silver controlled by each of the base-metal minerals but I doubt very much whether his method can be justified, chiefly because there appear to be more variables to be considered than the few that appear in his equations. Typical Examples of Irregular Silver Distribution Combination Mine, New Mexico, U.S.A.—In this mine an examination of the assay plans either of the mine or of the average mill feed or products

Jan 1, 1935

By Darwin Soler

This chapter will outline the operating procedures at Freeport Kaolin Co.5 kaolin mining operations in Georgia, with emphasis on explaining new techniques recently introduced to reclaim mined-out areas for reforestation. Kaolin is found in the central Georgia area (See Fig. 1) in lenticular, sedimentary deposits of high purity. The beds are Cretaceous in age and vary from 4 to 60 ft in thickness, under a cover of sedimentary overburden ranging from 5 to 150 ft. The kaolin beds contain the mineral kaolinite (a hydrous aluminum silicate) and have a particle-size distribution ranging from about 20µ to the submicron level and averaging about 60% below 2µ. The only contaminants in the deposits are coarse mica, quartz particles, and minor amounts of iron and titanium oxides. These kaolin deposits vary widely in quality from deposit to deposit and also vary within individual deposits. Some of the controlling quality specifications which are determined during exploration are viscosity in water suspension, optical brightness index as measured by reflectance, amenability to bleaching, and particle-size distribution. For this reason, only a small portion of the total kaolin reserves in the Georgia belt are suitable for wet processing for the present markets being served. Kaolin clays are selectively mined by different mechanical methods and are beneficiated by both wet and dry processing to produce a number of products used as fillers and coatings in the paper industry; fillers or additives in the paint, plastic, and rubber industries; and as basic raw materials for the ceramic industry. The State of Georgia is the largest kaolin producer in the United States, with 3,206,454 tons produced in 1966, according to U.S. Bureau of Mines statistics. Freeport Kaolin Co’s processing plant is located in Gordon, Ga., about 20 miles east of Macon, Ga.

Jan 1, 1969

By Harry V. Warren

During the past few years the author has had an opportunity to examine a number of base-metal mines in the western United States, France and Spain. Nearly all of these mines produced some silver, and laboratory studies of the relationship between the silver and the various base-metal minerals in the ore have given some interesting and, it is to be hoped, enlightening facts. As this work was undertaken primarily to ascertain the amount of silver chemically associated with, not controlled by, the various base-metal minerals, the laborious and not always satisfactory method of picking out "pure" samples of each mineral from many localities was employed. However, the results obtained have justified the time that has been spent. Not only have the conclusions arrived at in different mining areas been in accordance with one another but in many instances they have afforded scientific support or explanations for much of the knowledge that operators in the various camps have acquired by practical experience. Although many of the conclusions are not new, many of the facts are, and the results, if they do nothing else, serve to confirm the findings of some of the other workers in this field, particularly the work of Nissen and Hoyt1 and of F. N. Guild2. Dr. S. G. Lasky3 has suggested a novel method for determining from a series of algebraic equations the amount of silver controlled by each of the base-metal minerals but I doubt very much whether his method can be justified, chiefly because there appear to be more variables to be considered than the few that appear in his equations. Typical Examples of Irregular Silver Distribution Combination Mine, New Mexico, U.S.A.—In this mine an examination of the assay plans either of the mine or of the average mill feed or products

Jan 1, 1935

By Harry V. Warren

During the past few years the author has had an opportunity to examine a number of base-metal mines in the western United States, France and Spain. Nearly all of these mines produced some silver, and laboratory studies of the relationship between the silver and the various base-metal minerals in the ore have given some interesting and, it is to be hoped, enlightening facts. As this work was undertaken primarily to ascertain the amount of silver chemically associated with, not controlled by, the various base-metal minerals, the laborious and not always satisfactory method of picking out "pure" samples of each mineral from many localities was employed. However, the results obtained have justified the time that has been spent. Not only have the conclusions arrived at in different mining areas been in accordance with one another but in many instances they have afforded scientific support or explanations for much of the knowledge that operators in the various camps have acquired by practical experience. Although many of the conclusions are not new, many of the facts are, and the results, if they do nothing else, serve to confirm the findings of some of the other workers in this field, particularly the work of Nissen and Hoyt1 and of F. N. Guild2. Dr. S. G. Lasky3 has suggested a novel method for determining from a series of algebraic equations the amount of silver controlled by each of the base-metal minerals but I doubt very much whether his method can be justified, chiefly because there appear to be more variables to be considered than the few that appear in his equations. Typical Examples of Irregular Silver Distribution Combination Mine, New Mexico, U.S.A.—In this mine an examination of the assay plans either of the mine or of the average mill feed or products

Jan 1, 1935

During the past few years the author has had an opportunity to examine a number of base-metal mines in the western United States, France and Spain. Nearly all of these mines produced some silver, and laboratory studies of the relationship between the silver and the various base-metal minerals in the ore have given some interesting and, it is to be hoped, enlightening facts. As this work was undertaken primarily to ascertain the amount of silver chemically associated with, not controlled by, the various base-metal minerals, the laborious and not always satisfactory method of picking out "pure" samples of each mineral from many localities was employed. However, the results obtained have justified the time that has been spent. Not only have the conclusions arrived at in different mining areas been in accordance with one another but in many instances they have afforded scientific support or explanations for much of the knowledge that operators in the various camps have acquired by practical experience. Although many of the conclusions are not new, many of the facts are, and the results, if they do nothing else, serve to confirm the findings of some of the other workers in this field, particularly the work of Nissen and Hoyt1 and of F. N. Guild2. Dr. S. G. Lasky3 has suggested a novel method for determining from a series of algebraic equations the amount of silver controlled by each of the base-metal minerals but I doubt very much whether his method can be justified, chiefly because there appear to be more variables to be considered than the few that appear in his equations. Typical Examples of Irregular Silver Distribution Combination Mine, New Mexico, U.S.A.—In this mine an examination of the assay plans either of the mine or of the average mill feed or products

Jan 1, 1935

By A. J. Weinig

The method described is a modification of Schutzenberger's whereby the solution in titrations and the standards are protected from the atmosphere by a layer of kerosene. Indigo-disulfonate is the indicator and sodium hydrosulfite the standard. The method is rapid and delicate. It has been tried on a variety of mill solutions and mill men have found the method of value. THE important part that dissolved oxygen plays in the cyanide treat-ment of gold and silver ores is commonly recognized by most metallurgists and mill men.. But heretofore there has been no simple method on which the mill man could rely for determining the amount of oxygen in the various mill solutions. A few methods are adaptable to the, laboratory, such as the gasometric1 determination by actually measuring the oxygen. Winkler's2 iodometric method, and Schutzenberger's3 method, which depends on the reducing action of sodium hyposulfite on a blue solution of indigo carmine; but these are not readily adaptable to the needs of the mill man and cannot be applied to cyanide solutions without modification. The gasometric method is the most accurate but it takes too long to complete a determination. White's' method (a colorimetric method that depends on the degree of coloration imparted to a solution of pyrogallic acid in the presence of caustic soda) although much more rapid than any of the methods named, often gives misleading results on account of grading of colors of different solutions, which nearly always contain various kinds of salts. The method herein described was devised for the use of the mill man, the aim being to develop a simple, practical, accurate method. It is a modification of the Schutzenberger method and depends on the reducing

Jan 8, 1924

By D. L. Harris, A. M. Gaudin

Observations were made of the distribution of mercaptan containing S35 between aqueous solution and mineral and between aqueous solution and the gaseous phase. Although equilibrium may not have been attained, adsorption of the reagent was shown to occur reasily from air or aqueous solution on sphalerite, zincite, and willem-ite and to correspond to flotation. Adsorption on quartz did not similarly occur. THE following results, presented here in condensed form,' were obtained in a preliminary study of the adsorption of n-hexane thiol, hexyl mercaptan, on sphalerite, zincite, willemite, and quartz, from aqueous solution and from a gas. Interest in this subject was aroused by a Belgian report' of effective use of hexyl mercaptan for flotation collection of oxidized zinc minerals. The relatively low boiling point, 149°C, of the mercaptan3 suggested the desirability of extending the usual measurements of partition of collector between aqueous solution and gas and between gas and mineral. It is believed that this paper presents the first measurements of this type on a flotation system. Attempts were made to carry out the measurements at equilibrium, but as the work progressed it became increasingly doubtful that this desirable condition had been achieved. To control composition and extent of the gas phase, the apparatus was a wholly-enclosed thermally-controlled glass system. Because of these constraints and the desirability of dealing with pure minerals, a scale of operations was chosen in which a few grams of deslimed mineral were used in each test. It was also necessary to choose a particularly sensitive method for mercaptan analysis, and in fact a method that would permit the experimenters to follow the approach to equilibrium. For these reasons mercaptan marked by radiosulphur 35 was used. An analysis was made for the radiosulphur by a modification of the method of Gaudin and Carr. Coarsely-crystallized sphalerite was handpicked, stage-crushed in the dry state, wet-screened on a 200-mesh sieve, and deslimed in water at about 5 microns. Further treatment consisted of a wash in dilute aqueous hydrogen peroxide, drying, removal of the dark-colored fraction in a Frantz magnetic separator, washing in very dilute hydrochloric acid, repeated washing in distilled and conductivity water, and drying. The last washings showed a conductivity equivalent to a few ppm NaC1, that is, much more than would be provided, theoretically, by a saturated ZnS solution. The material was stored dry in sealed bottles. Analyses were as follows: Zn, 62.3 pct; Fe, 0.43 pct; Cd, 0.44 pct; S, 31.2 pct; Mn, 0.001 pct. The specific surface (BET method) was 2000 cm2/g. Zincite from Franklin furnace of the New Jersey Zinc Co. was hand-picked, dry-crushed, wet-screened at 100 mesh, and deslimed at about 10 microns. After drying, the associated zinc, manganese, calcium, and silicate minerals were removed in a Frantz magnetic separator. The purified zincite was washed in distilled water and conductivity water to a conductance of less than 2 ppm equivalent NaC1, dried, and stored. Analyses were as follows: Zn, 75.1 pct; Fe, 0.9 pct; Mn, 2.78 pct. The specific surface (BET method) was 1740 cm 2/g. Willemite, also from Franklin furnace, was purified similarly. Analyses were as follows: Zn, 52.5 pct; Fe, 0.12 pct; SiO², 27.3 pct; loss on ignition, 0.13 pct. The specific surface was 1760 cm 2/g. Conductivity water (double-distilled) and demin-eralized-distilled water were used in most of the tests. The specific resistance was not less than 600, 000 ohms, and usually above 1,000,000. Radiosulphur-marked hexyl mercaptan (1-hexane thiol) was synthesized by Tracerlab, Inc., Boston. Two lots were secured several months apart. The last lot, consisting of about 0.5 g of the mercaptan, had a total activity of about 10 millicuries. Tracerlab Co. guaranteed only the activity; hence a quasi -vapor pressure determination (based upon an S analysis) of the mercaptan was made. The calculated value, 4.2 mm of mercury at 25.5' C, has been compared with that of a sample of Highest Purity 1-hexane thiol from Fisher Scientific Co. The latter had a vapor pressure of 4.5 mm of mercury at 2.5 C. Analytical Procedures The sample containing radiosulphur-marked mercaptan was oxidized to convert the mercaptan sulphur to sulphate, carrier barium sulphate being added to provide a suitable quantity of total barium sulphate in a filter cake. The precipitate was filtered and dried, and counting was carried out either in a streaming-gas (Q-gas) counter for high sensitivity or with an end-window G-M counter for convenience. The oxidized and precipitated mercaptan gave a radioactive count of 65 counts per minute per microgram in the end-window Geiger-Mueller counter and 1100 counts per minute per microgram in a Q-gas counter. For standardization of the mercaptan solution, 15 replicate analyses were made. The average deviation per measurement was about 1600 cpm in 65,000 cpm, the probable error in the mean being 275 cpm. It

Jan 1, 1955

By K. F. Veit, R. L. Marovelli, T. S. Chen

There is an error at the bottom of the right-hand column of p. 11, and the top left-hand column of p. 12. The first paragraph of the section EXPERIMENTAL RESULTS should read: "Thermal shock tests were conducted on circular discs of basalt, quartzite, and taconite. Average grain sizes ranged between 0.10 and 0.25 mm. Individual discs were tested only once, either in a cooling or heating shock. Several discs of the same rock type were shocked at a temperature, and the temperature level then was increased for another test by 10° to 20°F until complete fracture occurred when a specimen was quenched. The minimum temperature difference required to initiate cracking, TmiN was observed and used, together with other property values, to compute the dimensionless stress o*max . The observed ?Tmin for cooling shock in different media is tabulated in Table I. The property values E, St, a, and K are shown in Figs. 6, 8, 10, and 11, respectively. Discs of various rocks shocked in different isothermal baths gave o*max values for a wide range of Biot numbers. These o*max values were then compared with the theoretically predicted values."

Jan 1, 1968

By A. H. Larson, R. J. McClincy

Laboratory-scale decopperizing experiments with multiple sulfur addifions were conducted at 330°C on ternary Pb-Cu alloys containing, as the third elenlent, Sn, Ag, As, Sb, Bi, Zn, and Au, common impurities in lead blast-furnace bullion. For silver and tin, an increased rate and extent of 'cofifier removal was obsert3ed. The elements As, Sb, Zn, Au, and Bi had no effect or less effect as compared to sulfur additions with no i)npurily additions. THE production of primary lead in the blast furnace yields an impure lead frequently containing such impurities as copper. antimony. arsenic. tin, gold, silver iron, oxygen. and sulfur. By cooling this lead to a temperature near its melting point. most of the iron, sulfur, and oxygen and part of the other impurities are removed in the form of a dross. With incipient solidification of the lead, the copper concentration wil have been reduced to 0.02 to 0.05 pct. depending upon the concentration of the other impurities. according to Davey.' Since copper interferes with the treatment of silver after the desilverizing process, it is desirable to decrease the copper content of the lead still fur-ther before the lead is desilvered. The decopperizing of the lead is accomplished by stirring a small quantity. approximately 0.1 pct. of elemental sulfur into the lead at a temperature near its melting point, 330" to 360°C. The copper is removed as a copper sulfide which constitutes a small fraction of a voluminous dross consisting mostly of lead sulfide and entrained metallic lead. The residual copper concentration following the decopperizing operation is frequently as low as 0.001 to 0.005 pct. Thi fact has aroused considerable interest because the equilibrium copper concentration of lead in contact with solid PbS and solid Cu2S is at least an order of magnitude greater, 0.05 pct Cu at 330C. 1, 2 Most investigators have suggested that various impurities in the lead bullion are responsible for the very low copper concentrations frequently encountered in practice. There is little agreement, however? as to which of the impurities are helpful and which are not.3"11 Also. few investigators have sought to explain the mechanisms responsible for the removal of copper to very low concentrations. Willis and Blanks9 have proposed that a nonstoichiometric copper-deficient cuprous sulfide forms in place of the supposed Cu2S. Being copper-deficient, this sulfide phase would possess a low copper activity, and the diffusion of copper dissolved in the liquid lead into this phase would be greatly facilitated. Pin and wagner2 have investigated the removal of copper from liquid lead by studying the effect of impurity-doped lead sulfide on the decopperizing of pure Pb-Cu alloys. Samples of the doped PbS were held in contact with copper-saturated lead for 1 week at 33'7°C. They reported a beneficial effect on decopperizing with bismuth and antimony and no effect with tin or silver. which is directly opposite to the results observed in practice and those reported by Davey 3 and this studv. The purpose of this paper is to describe the effects of certain additive elements on the extent to which copper can be removed fro111 liquid lead by successive additions of sulfur. The impurity elements were added individually to prepared Pb-Cu alloys. The resulting ternary alloys as well as a binary Pb-Cu alloy were then decopperized with repeated additions of sulfur. EXPERIMENTAL Materials. Granulated test lead with a purity of 99.999 pct and the additive elements Cu. Ag. Sb. Bi. Zn. Sn. and Au with purities of 99.99 pct were American Smelting and Refining Co. research-grade materials. The major impurities in the lead were 1 ppm each of iron and copper. all others being less than 1 ppm. The arsenic used was a technical-grade arsenic of 98+ pct purity. Reagent-grade flowers of sulfur were melted under argon to provide small pieces free of fines. Apparatus. The decopperizing experiments were carried out in a 25-mm-OD by 375-mm-long Pyrex tube sealed at one end. The tube was mounted vertically in a resistance-heated. hinge-type tube furnace controlled to within ±lcC. Temperature measurement was accomplished by means of a standardized chromel-alumel thermocouple sealed into the base of a silica. paddle-type stirring rod. All decopperizing experiments were carried out under an argon atmosphere. Procedure. A Pb-Cu starting alloy containing 0.05 pet Cu was prepared under carbon and poured into cold tap water to produce shot. The ternary alloys were prepared by melting together 100 g of the starting alloy and a sufficient amount of the impurity element to yield the desired concentration. The resulting alloy was then homogenized in a Pyrex tube at 450C with continuous stirring. The furnace temperature was then lowered to the operating temperature of 330°C. When thermal equilibrium had been obtained at the operating temperature, individual additions of 0.2 pct (0.2 g) of solid sulfur were added to the melt and stirred in. Stirring was continued for a period of 3 min. discontinued for 5 min. and resumed for the remaining 2 min of a 10-min cycle. This cycle was repeated for as many sulfur additions as desired. When the decopperizing experiment had been completed the lead bullion was quenched and samples of the bullion and dross phases were taken for analysis. Results. The results obtained in the decopperizing

Jan 1, 1970

By L. H. Schwartz, S. K. Lahiri, M. E. Fine, D. Chandra

WHILE the yield stress of solution treated Fe-Cu alloys increases rapidly with aging, a precipitate has only been directly observed in overaged samples.'-" This precipitate is essentially pure fcc copper, the E phase. Hornbogen2 suggested that the initial hardening is due to clustering of copper atoms in a bcc iron matrix. Hornbogen was not able to obtain direct confirmation for such bcc clusters by either X-ray diffraction or electron microscopy; however, he did observe analogous clustering in Fe-Au alloys7 where the scattering factor difference between iron and gold is more favorable than that for iron and copper. Fujii, Nemoto, Suto, and Monma4 recently observed diffuse electron diffraction scattering in a 2.5 pct Cu alloy aged at 500°C: for 30 hr. They suggested that part of the diffuse SCattering was due to a modulated structure of 26 to 42 wavelength parallel to the (110) matrix plane. This disappeared with further aging. However, 30 hr at 500°C is well beyond the peak strengthening in this alloy which occurred at less than I hr at 500oc.4 By transmission electron microscopy they could not distinguish between quenched samples and samples aged to maximum strength. The purpose of the present investigation was to try to obtain further confirming evidence for a precipitation stage preceding the formation of the fcc E phase. Change in Young's modulus was selected for study be- EXPERIMENTAL PROCEDURE The Fe-1.67 at. pct Cu alloy was supplied by the U.S. Steel Corporation in the form of a hot forged rectangular bar. The carbon and nitrogen contents of this alloy were 70 ppm and less than 10 ppm, respectively. The samples for Young's modulus measurements were made by cold swaging to & in. diam and cutting to 2 in. lengths. Tensile test samples 1 in. long were prepared by cold rolling to a thickness of 0.012 in. and then machining to gage dimensions of 0.5 by 0.012 in. For Mijssbauer studies, foils 1 mil thick were prepared by cold rolling. During fabrication the alloy was given intermediate anneals at 845OC followed by quenching. The specimens after preparation were solution treated at 1000°C (y field) or 845°C (a field) for 1 to 2 and 5 to 6 hr, respectively, in a dynamic vacuum of about 2 x lom6 torr. The tensile and modulus specimens were drop-quenched into iced brine through an aluminum foil which vacuum-sealed the bottom of the furnace tube. The MOssbauer foils were drop-quenched into helium gas cooled by passing through N2 (liq) in a copper coil. The grain diameter of the modulus and tensile specimens was between 0.025 and 0.035 mm. The samples were aged from 400" to 500°C in a dynamic vacuum of 2 x 10-6 torr and cooled by pulling into a water-cooled portion of the vacuum system. The dynamic Young's modulus was measured at room temperature using electrostatic excitation and detection of longitudinal vibrations.' Neglecting any length change during the aging, (this is expected to be comparatively small), the fractional change of Young's modulus (?Et)/Eav, is equal to twice the fractional change of the resonant frequency (?ft)/fav . ?ft =ft —fo where fo and ft are the resonant frequencies of a specimen before aging and after aging to time t. The measured f's, approximately 51,000 cps, were corrected to 25°C for any variation in the ambient temperature. The tensile tests reported were carried out at room temperature in an Instron with a 1000 lb full scale load cell. The flow stress reported herein corresponds to a strain of 0.002. The Massbauer spectra were taken in a constant acceleration model AM-1 NSEC spectrometer. The

Jan 1, 1970

By V. Koump, T. F. Perzak, R. H. Tien

A series of experiments were carried out in which the rates of propagation of the liquidus and the eutectic fronts Mere measured during essentially one-dimensional freezing of Al-Cu alloys. The dimensions of the ingots were 3 by 5 by 6 in. Three different alloys containing 0.1, 4.5, and 17 pct Cu were used in these experitments. For each alloy the rate of heat removal was varied to give a total jreezing time in the range 3 to 30 min. The results of these measurements cowlpared favorably with the theoretical model of freezing of binary alloys with time-dependent surface temperature. IN engineering analysis of solidification of commercia1 steels and nonferrous alloys it is a common practice to assume that an alloy freezes by propagation of an isothermal solidification front, i.e., essentially as a pure metal. In two recent theoretical investigations'j2 the present authors explored the possibility of a more realistic approach to the problem of solidification of alloys. In the proposed model the freezing of an alloy is assumed to take place by propagation of two isothermal fronts, i.e., the liquidus front and the solidus (or eutectic) front. The region between the two fronts contains both liquid and solid and is referred to as the solid-liquid region. The width and the solid content of the solid-liquid region vary with alloy type, solute concentration, and cooling rate. For a given alloy system, initial concentration of solute, and the mode of heat removal, the proposed model yields the temperature distribution within the solid skin, temperature, solid fraction, and concentration distributions with the solid-liquid region, and the rates of propagation of the liquidus and the solidus fronts. This model is obviously of considerable practical importance in engineering analysis of solidification processes, since it gives a more realistic estimate of skin strength during solidification and a better estimate of the total freezing time. Before the new model can be used with confidence, however, it is necessary to test this model experimentally. The experimental testing of the proposed model is a relatively simple matter since the effects to be measured are large and a relatively simple experiment will suffice. The theoretical model predicts, for example, that during freezing of an alloy containing substitutional type solute (negligible diffusion in the solid during freezing) the solid-liquid region occupies an appreciable portion of the ingot, even at low concentration of solute.' Another prediction of the theo- V. KOUMP, formerly with U. S. Steel Corp., is now with Research and Development Center, Systems and Process Division, Westinghouse Electric Corp., Pittsburgh, Pa. R. H. TlEN is Senior Scientist, Fundamental Research Laboratory, U. S. Steel Corp., Research Center, Monroe ville, Pa. T. F. PERZAK, formerly with U.S. Steel Corp., is now with Fiber Industries, Greenville, S. C. Manuscript submitted March 6, 1968. IMD retical model, easily verifiable by experiment, is that the rate of propagation of the solidus (or eutectic) front increases as the solidus front approaches the center of the slab. This prediction is contrary to well-known behavior of the solidification front during freezing of pure metals, where the rate of propagation of the solidification front decreases with time and freezing is completed at the lowest rate. A rather severe test of the proposed model is provided by comparison of theoretical predictions and experimental measurements of the effects of cooling rate and composition on the rates of propagation of the liquidus and the eutectic fronts. In order to test the soundness of the formulation and the method of solution of the problem of solidification of alloys a series of experiments were carried out in which the rates of propagation of the liquidus and the eutectic fronts were measured during essentially one-dimensional solidification of A1-Cu alloys. The A1-Cu system was chosen strictly as a matter of convenience. Three different alloys containing 0.1, 4.5, and 17 pct Cu were used in these experiments. For each alloy the rate of heat removal was varied to give the total freezing time in the range 3 to 30 min. The results of these measurements are compared with the predictions of the theoretical model of solidification of binary alloys, with time-dependent surface temperature.' Before the experiments described in this paper were undertaken, a serious attempt was made to utilize the measurements of previous investigators to test the theoretical model. In the course of this preliminary study a careful review was made of experiments of Pellini and coworkers3 and Doherty and Melf~rd.~ The measurements in Pellini's work were carried out using a steel containing at least four major components. Evaluation of the solid fraction-temperature relation for this steel (required in the theoretical model) is difficult and uncertain. Doherty and Melford, on the other hand, measured the solid fraction-temperature relation experimentally, but did not give sufficient data to explore the effects of composition and the cooling rates on solidification. Hence it was not possible to utilize these measurements to test our theoretical model. EXPERIMENTAL METHOD The experimental technique used in this investigation differs somewhat from the more conventional techniques employed in solidification studies. This technique was developed primarily to eliminate con-vective mixing in the molten metal caused by pouring of molten metal into the mold. In our experiments A1-Cu alloys were melted directly in the mold. The mold assembly used in solidification experiments is shown in Fig. 1. The mold was fabricated from *-in. stainless-steel sheet. The dimensions of

Jan 1, 1969

By S. Chander, D. W. Fuerstenau

The inherently sluggish nature of the reactions involving sulfide minerals, particularly molybdenite, in aqueous solutions under ambient flotation conditions make their investigation complicated and difficult. Since the physicochemical nature of sulfides is known to be controlled by electrochemical reactions, a number of attempts have been made in the past to study sulfide mineral-collector interactions through rest potential and electrode polarization measurements. The purpose of the present investigation was to study the interaction of potassium diethyldithiophosphate (KDTP) with molybdenite. In the present investigation of the interaction of KDTP at the molydenite surface, molybdenite electrode rest potential measurements and electrode polarization behavior were experimentally determined under various solution conditions. The standard reduction potential for the equilibrium between dithiophosphate ions (DTP-) and bis(diethoxythiophosphoryl)disulfide (DTP) [(DTP). + 2e :i7-± 2DTP- (1)] can be evaluated from the electrode rest potential measurements on a platinum electrode in deoxygenated KDTP solutions saturated with freshly prepared (DTP)2. Experimental A platinum wire was used as the electrode for potential measurements. Prior to each measurement the elec-

Jan 1, 1975

By H. W. Imholz

Considerable stimulation was given to the interest in the lower producing zones of the North Central Texas area when several wildcat wells developed production in these formations. The Iron Mountain and Humble-Olander well in sec. 196, B.B.B. and C. survey, Jones County, Texas, had a potential of 7500 bbl. (after a 3000-gal. acid treatment) from the Palo Pinto limestone encountered from 3238 to 3250 ft. This well heightened the prospects for oil in commercial quantities in the Canyon, Strawn, Bend and Ellenburger formations in the adjoining counties of Shackelford, Haskell, Throckmorton and others. The Hickok-Van Parmer well in sec. 477 S.P.R.R. Co. survey, Eastland County, Texas, was good for 1040 bbl., natural, from the Ellen-burger limestone, encountered from 4024 to 4033 feet. The Hal Hughes-Cook and Jordan (Poindexter) well in sec. 350, block 6, S.P.R.R. Co. survey, Callahan County, flowed 890 bbl. in 9 hr., after acid treatment, from the Ellenburger limestone, encountered from 4369 to 4388 feet. Both of the Ellenburger wells apparently are maintaining the pressure and volume of their original potential tests, and will no doubt open up considerable area in the adjoining counties for prospecting to the Ellen-burger limestone. Production increased in the district from 12,081,871 bbl. in 1936 to 15,384,500 bbl. The increase was due to developments in Jones and Jack Counties. Jack County production increased from 896,680 bbl. in 1936 to 3,117,063 in 1937, while Jones County increased from 1,596,743 bbl. in 1936 to 2,289,322 in 1937. The remaining counties of the district showed a slight decrease in the amount of oil produced.

Jan 1, 1938

By M. D. Everell

A study has been made of the relationships between grinding selection functions and physical properties of rocks, particularly the failure loads of irregular particles and the compressive and tensile strengths of large regular specimens. The rocks used were limestones of various structures and impurity contents, a carbonatite, and a high-purity crystallized calcite. The selection functions were obtained from carefully performed grinding experiments on selected narrow size ranges in various grinding machines-a laboratory batch ball mill, a Hard grove grindability machine, and a mill of the type suggested by Yang, Mempel, and Fuerstenau. The specific rates of breakage during milling and the failure loads of irregular specimens correlated well for the rocks studied. The testing of irregular particles is suggested as a means of predicting the behavior of commercial mills.

Jan 1, 1973

By E. Cortese

Phosphate occurs in many places in Egypt, in two main zones: one in Upper Egypt, along the Nile Valley, principally on the right side, and one near the Red Sea coast. In the Nile zone, the principal layers of phosphate rock are to be found: in the mountains east of Mahamid (right side of the valley); in the hills near Sibaiya, also called Sebaieh (both sides); in the hills of Agula, Hagaza and Gurn (right side); and further on, along the camel road leading from Kena to Kosseir, a few miles east of Legheita's Oasis. Following that track, after having crossed the watershed, underlain by rocks like granite, serpentine, porphyry, and diabase, one comes again to the Cretaceous formation containing considerable quantities of phosphate, at Gebel Duwi, Gebel Nakheil, and Gebel Kosseir El Kadim, not far from the coast. More to the north, in the hills near Safaga Bay (Um El Uettat El Wasif) is another rich bed of the same formation. There are reports of phosphate existing near Suez and in the Sinai peninsula, and possibly some other zones can be traced or discovered in the future. The mines are well developed on the right bank of the Nile at Sharawna (near Sebaieh Station) and along the Red Sea coast, at Safaga and near Kosseir. Prospecting work has been done at Gebel Duwi, Gebel Gurn, Agula, Hagaza, and near Mahamid. The phosphatic formation in Egypt is quite variable, not only from place to place, but even within small areas. The phosphate is all of Cretaceous age, as is proved by the fossils found in it. It is considered to belong to the Campanian or, more appropriately, to the Maestrichtian series. The principal fossils are: Alectryonia Villei Coq. sp.; Roudereia Aures-sensis Coq. sp. = R. Drui Mun. Chalm.; Trigonoarca multidentata, Newton.; Cardita libyca, Zittel; Libycoceras Ismaely, Zittel; Nautilus sp.; Ancystrodon libycum Dam.; Corax pristodontus Ag.; Lamma biauriculata Zittel sp. Above the Maestrichtian beds, we have the Danian series, unconform-ably covered by Lower Eocene beds. Under the phosphate is a grayish marly clay, and beneath this, bluish-gray scaly shales resting on the Nubian sandstone (Santonian).

Jan 1, 1918

By Gerald A. Munson, Fremont F. Clarke

DURING recent years the use of lithium has expanded greatly in industrial, chemical, and metallurgical fields, while at the same time modernized methods of mining and refining lithium have increased production. Technical literature includes many papers describing the geology and mineralogy of lithium deposits. Mining and beneficiating problems, however, have not been thoroughly described. Published reports have failed to emphasize that one of the chief reasons lithium minerals were not extensively mined until recently is the difficulty of beneficiation. Four lithium minerals of pegmatites and the lithium-sodium-phosphate byproduct from the brines at Searles Lake, Calif., have been sources of lithium. Among the pegmatite minerals, only spodu-mene and petaline are known to occur in deposits large enough to support large tonnage operations. Spodumene, a lithium-aluminum-silicate, is the principal lithium mineral mined in the U. S. Lithium Corp. of America, during its rapid development and growth, has successfully employed three different methods of concentration: 1) hand sorting, 2) Heavy Media separation, and 3) froth flotation. Each method, though economic in its sphere, is becoming outdated as extractive techniques improve. The most recent development has taken place at Lithium Corp.'s new operation in North Carolina, where the extractive process yields lithium compounds directly from mined rock without an intervening concentrating step. Until 1954 Lithium Corp. mined only in the southern part of the Black Hills of South Dakota, where the various types of lithium-bearing pegmatite could be used to develop untried processes of beneficiation. Concentrates from these operations are shipped to the corporation's chemical plant in St. Louis Park, Minn., where actual extraction of lithium and production of commercial compounds takes place. Nature of Deposits The geology of the southern Black Hills pegmatite deposits is admirably presented by a U. S. Geological Survey publication.' Only the factors pertinent to exploitation need be mentioned here. Lithium Corp. has been interested primarily in three deposits, the Edison, the Mateen, and the Longview-Beecher No. 2. The history of these deposits dates back to the mining boom of 1880 to 1900 in the Black Hills, when most of the deposits were prospected for tin by the ill-fated Harney Peak Tin Mining Co. The three deposits described here, however, lay idle and unwanted until Lithium Corp. opened the Edison mine in 1943 and the Mateen and Longview-Beecher No. 2 in 1951. Edison Deposit: The Edison deposit is about a mile southeast of Keystone, S. D., on the northeastern flank of the Harney Peak uplift. The deposit consists of no less than four and probably at least six separate pegmatites that coalesce in a central mass from which the individual pegmatites finger outward and downward in complexly folded schist and gneiss beds of the country rock.

Jan 1, 1956

By H. OSBORG

THE patent literature of alloys for the last two decades or so indicates that the number of liatents referring to smaller and smaller percentages of essential alloying constituents is on the increase, whereas the number of patents relating to considerable percentages of alloying additions becomes smaller. The claims regarding physical properties indicate consistently higher and more rigid values and show the possibility of modifying the physical properties by only slight changes in the percentage of minor constituents or by applying different methods - of treatment. Alloys of the same base have been made suitable for a number of different applications by intelligently and rigid- ly controlling their composition, melting, casting, working, and heat-treatments. Other trends indicated in the development of modern alloys are higher corrosion resistance, higher fatigue limits, and "stability" in volume.

Jan 1, 1937

By Takashi Katsura, Avnulf Muan

Equilibrium relations in the system FeO-Fe2O3 Cr2O3 have been determined at 1300°C at oxygen pressures ranging from that of air (0.21 atm) to 1.5 x 10-11 atm. The following oxide phases have stable equilibrium existence under these conditions : a sesquioxide solid solution with corundum-type structure (approximate composition Fe2O3-Cr2O3); a ternary solid solution with spinel-type structure (approximate composition FeO Fe2O3-FeO Cr2O3) and a ternary wüstite solid solution with periclase-type structure and compositions approaching FeO. The metal phase occurring in equilibrium with oxide phase(s) at the lowest oxygen pressures used in the present investigation is almost pure iron. The extent of solid-solution areas and the location of oxygen isobars have been determined. ThE system Fe-Cr-O has attracted a great deal of interest among metallurgists as well as ceramists and geochemists. Metallurgists have studied the system because of its importance in deoxidation equilibria, ceramists because of its importance in basic brick technology, and geochemists because of its importance for an understanding of natural chromite deposits. Chen and chipman1 investigated the Cr-O equilibrium in liquid iron at 1595°C in atmospheres of known oxygen pressures (controlled H2O/H2 ratios). The main purpose of their work was to determine the stability range of the iron-chromite phase. Hilty et al.2 studied oxide phases in equilibrium with liquid Fe-Cr alloys at 1550°, 1600°, and 1650°C. They reported the existence of two previously unknown oxide phases, one a distorted spinel with composition intermediate between FeO Cr203 and Cr3O4, the other Cr3O4 with tetragonal structure. They also sketched diagrams showing the inferred liqui-dus surface and the inferred 1600°C isothermal section for the system Fe-Cr-O. Koch et al3 studied oxide inclusions in Fe-Cr alloys and also observed the distorted spinel phase reported by Hilty et al. Richards and white4 as well as Woodhouse and White5 investigated spinel-sesquioxide equilibria in the system Fe-Cr-O in air in the temperature range of 1420" to 1650°C, and Muan and Somiya6 delineated approximate phase relations in the system in air from 1400" to 2050°C. The present study was carried out at a constant temperature of 1300° C and at oxygen pressures ranging from 0.21 atm (air) to 1.5 x 10-11 atm. The chosen temperature is high enough to permit equilibrium to be attained within a reasonable period of time within most composition areas of the system, and still low enough to permit use of experimental methods which give highly accurate and reliable results. These methods are described in detail in the following. I) EXPERIMENTAL METHODS 1) General Procedures. Two different experimental methods were used in the present investigation: quenching and thermogravimetry. In the quenching method, oxide samples were heated at chosen temperature and chosen oxygen pressure until equilibrium was attained among gas and condensed phases. The samples were then quenched rapidly to room temperature and the phases present determined by X-ray and microscopic examination. Total compositions were determined by chemical analysis after quenching. In the thermogravimetric method, pellets of oxide mixtures were suspended by a thin platinum wire from one beam of an analytical balance, and the weight changes were recorded as a function of oxygen pressure at constant temperature. The data thus obtained were used to locate oxygen isobars. The courses of the latter curves reflect changes in phase assemblages and serve to supplement the observations made by the quenching technique. 2) Materials. Analytical-grade Fe2O3 and Cr2O3 were used as starting materials. Each oxide was first heated separately in air at 1000°C for several hours. Mixtures of desired ratios of the two oxides were then prepared. Each mixture was finely ground and mixed, and heated at 1250" to 1300°C in air for 2 hr, ground and mixed again and heated at the same temperature for 5 to 24 hr, depending on the Cr2O3 content of the mixture. A homogeneous sesquioxide solid solution between the two end members resulted from this treatment. A Part of some of the sesquioxide samples thus prepared was heated for 2 to 3 hr at 1300°C and oxygen pressures of 10-7 or 1.5 x 10-11 atm. Reduced samples (either iron chromite

Jan 1, 1964