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By D. A. Bailey, F. L. Kadey

During recent years, expanded perlite has found extensive use in a wide variety of commercial applications. Specifications for a range of uses call for aggregates having different physical properties and requiring various types of processing. The mitre scopic examination of expanded aggregates in specially prepared thin sections has yielded important information regarding internal structure and can be a useful aid in evaluation of the product. Perlite is an acid volcanic glass containing from 2 to 5 pct combined water. When crushed and rapidly raised to a temperature of 1500° to 2100°F, the volatilization of the combined water, coincident with softening of the siliceous glass, causes the particles to expand or pop. Expansion of perlite to densities as low as 1-1/2 lb per cu ft has been achieved, although the normal range is from about 2.5 to 10 lb per cu ft. Low-density materials of this nature have found application in cryogenic insula- tions, insulating tile and board, plaster and lightweight concrete, soil conditioners, filter aids after comminution, and a host of other uses. The expansion of perlite may be controlled by a number of variables. Some of these are ore milling, furnace design, expanding temperature, use of preheating techniques, screen size distribution, and feed rate of the crude. For the specific intended uses, the art of perlite expanding has developed to the point that much is known about the effect of these variables on the end product. The requirements, however, for an aggregate to be used in plaster are different from those for an expanded perlite to be used in lightweight concrete or for cryogenics. In the study of perlite, many practical questions remain unanswered. One of the most important is, "why do some perlites make hard aggregates while others cannot resist crushing or compaction?" The usual tests and methods for evaluation of nonmetallic substances have been tried and have contributed their share to our knowledge. But chemical analyses, compaction resistance tests, screen analysis, etc., will reveal only a limited amount of information. The answers to these complex questions may well be found in an examination of the intemal structure of the expanded particle. Such questions and the ex-

Jan 1, 1962

By P. Campbell, E. M. Gardy, L. Hachkowski

Introduction The efficient process control of a uranium mill and its associated mining operation requires a nearly continuous knowledge of the uranium concentration in the ore. Generally, the approach is to use laboratory assays of grab samples from the mill feed belt. In some cases, elaborate and expensive systems have been used to ensure random sampling, but even with these systems, mass-balance discrepancies still exist. There is a requirement for an on-line instrument that can give a prompt, accurate analysis of a large portion of the feed stock. The authors have recently evaluated a laboratory system that achieves these goals using neutron activation and delayed neutron counting. The development of the on-line uranium ore analyzer is a consequence of previous work done at the Whiteshell Nuclear Research Establishment (WNRE) (Campbell et al., 1978 and 1981), and is based on the emission of delayed neutrons after the neutron irradiation of fissionable materials. The mechanism of delayed neutron emission has been described (Keepin, 1965), but briefly is as follows. The fission fragments resulting from irradiation are in an excited state. Certain of these delayed fission products, precursors, decay to a more stable state by the emission of a delayed neutron. The delayed neutrons can be divided into six groups with effective half-lives ranging from 0.2-55 sec. It is important to note that more than 50% of the delayed neutrons are emitted within the first 6 sec after irradiation; this has an influence on the design of the analysis instrument. Measurement Cycle The measurement cycle involves a neutron irradiation, a delay time, and a counting period. The laboratory demonstration model of the ore analyzer comprises a californium-252 neutron source, a source chamber, a pneumatic transfer facility, and a neutron irradiation and counting module [(Fig. 1)]. The ore sample is placed in the irradiation chamber. The neutron source is transferred to the center of the chamber for a 5-sec irradiation while the helium-3 thermal neutron detectors are switched off. As the source returns to the storage chamber, the detectors are switched on and the delayed neutrons counted. The total assay cycle is 16 sec. The current instrument design calls for the ore to be cut from the main belt every 16 sec by a mechanical arm and dropped into the analyzer directly, or via another belt. The cycle starts when there is a sufficient mass of ore in the irradiation chamber to depress a switch and stops when the delayed neutron count is finished. The ore is then dropped through the bottom of the analyzer chamber and returned to the main belt. The counts from each cycle are accumulated and the data reduced to give the desired information.

Jan 1, 1985

By A. J. R. Curtis

THE Portland Cement Association was organized more than a third of a century ago by a group of cement manufacturers, to do cooperatively the educational and research work needed to ensure proper use of their product. Specifically, the Association has a threefold purpose: (1) to search out facts to ensure the best possible service from concrete; (2) to give the widest possible circulation to facts affecting the use of concrete; (3) to urge practical application of those facts on actual building projects. An organization of technically trained men conducts many research projects in the Association laboratories, interprets their findings and applies them practically in the field. The net effect of this work has been to improve the design and prolong the life of roads, bridges, industrial, commercial and public buildings, homes, farm structures and concrete products. The results of this technical research and experimentation are made available to the entire construction industry through bulletins and reports. Information is also released to the general public through newspapers, magazines and Association publications. In addition to a technical staff in the general office in Chicago, the Association has district offices in 22 cities. Field engineers cover 37 states, the District of Columbia and British Columbia, to give service to users of Portland cement. ACCIDENT PREVENTION During the early part of the century, there began a complete transformation in the processes of manufacturing cement. Electricity, with its attendant hazards, was beginning to replace steam; huge rotary kilns and tube mills were being substituted for the vertical kiln and "coffee grinder" mill. Only a few far-sighted leaders realized the import of these innovations, and they also recognized the fact that workmen must be protected against the hazards of the new machinery. By their sincerity and firm purpose they infused into the whole industry the concept of industry's responsibility for safeguarding employees. Even before effective compensation laws had been passed in most states, the Portland Cement Association had assigned to an Accident

Jan 1, 1937

By D. A Dahlstrom

THE necessity for slime elimination from valuable mineral and coal products has become increasingly significant within the past 5 years.' , Most of the mechanized mining and present beneficiation methods have resulted in larger tonnages of extreme fines, silts, and clay, commonly called slimes, processed by the preparation plant. Economics in turn have forced the plant operator constantly to lower the bottom size of effective beneficiation, necessitating an almost complete recovery of valuable solids regardless of particle size. When these important factors are considered, the need for an efficient slime removal method for fine solids becomes self-evident. Ordinarily slimes have been classified as ranging from 75 microns, or 200 mesh, down to colloidal dimensions of less than one micron. At the same time there is general agreement that the most troublesome slimes are primarily clays, silts and extreme fines, so that the actual top size may be placed at approximately 30 to 40 microns. This immediately restricts any desliming operation to classification, as screening or other methods would be both uneconomical and grossly inefficient at such dimensions. The liquid-solid cyclone has been recognized as one of the most efficient fine size classifiers available to the mining industry. Applications of cyclones to thickening, high efficiency solid recovery, degritting,

Jan 1, 1953

By D. A. Dahlstrom

THE necessity for slime elimination from valuable mineral and coal products has become increasingly significant within the past 5 years.' , Most of the mechanized mining and present beneficiation methods have resulted in larger tonnages of extreme fines, silts, and clay, commonly called slimes, processed by the preparation plant. Economics in turn have forced the plant operator constantly to lower the bottom size of effective beneficiation, necessitating an almost complete recovery of valuable solids regardless of particle size. When these important factors are considered, the need for an efficient slime removal method for fine solids becomes self-evident. Ordinarily slimes have been classified as ranging from 75 microns, or 200 mesh, down to colloidal dimensions of less than one micron. At the same time there is general agreement that the most troublesome slimes are primarily clays, silts and extreme fines, so that the actual top size may be placed at approximately 30 to 40 microns. This immediately restricts any desliming operation to classification, as screening or other methods would be both uneconomical and grossly inefficient at such dimensions. The liquid-solid cyclone has been recognized as one of the most efficient fine size classifiers available to the mining industry. Applications of cyclones to thickening, high efficiency solid recovery, degritting,

Jan 1, 1953

By A. M. Kivari

OPERATIONS at the cement plant of the Permanente Corporation, in the hills about 45 miles south of San Francisco and 12 miles west of San Jose, are interesting to the members because of the adoption of modern hydrometallurgical methods for handling and beneficiating raw materials. Closed-circuit grinding, classification, hydro-separation, flotation, agitation, and thickening are familiar terms at this cement plant. These methods make it possible to process refractory limestone into suitable materials for the manufacture of modern cements. Construction of the plant, having originally a design capacity of 7000 bbl. of cement per day, was started in June 1939, and the first clinker was ground on the following Christmas Day. At present the capacity is about 12,000 bbl. of finished cement per 24 hr., most of the production being low-heat cement.† About 3130 tons of limestone, 620 tons 0f clay and 100 tons of gypsum are required daily for this output. Occasionally iron oxide is added to the clay to obtain the desired composition. This, together with the gypsum, is purchased from outside sources. Up to the present time, the corporation has produced low-heat, normal, moderate heat, moderate-sulphate-resisting, high-early-strength, oil-well, modified-oil-well, and plastic cements. Most of the present production is for Shasta Dam, near Redding, Calif., the total demand being 5,800,000 bbl, of cement. Following are some of the cement requirements of the Shasta Dam specifications: [PERCENT Chemical composition: Loss in ignition2.25 Insoluble residue0.75 SO32.25 MgO5.00 Uncombined lime in clinker (CaO)1 .25 Percentage of iron oxide to alumina ratio1.5 Compound composition: Tricalcium silicate (3CaO.Si02)Not more than 3,5 Dicalcium silicate (2CaO.Si02)Not less than 40 and not more than 65 Tricalcium aluminate (3CaO.Al2O7)Not more than 7 Tetracalcium aluminoferrite (4CaO.A12O3.Fe203) Not more than 0 Fineness:] The average specific surface in the bin sampled shall be not less than 1800 sq. cm. per gram and no individual sample less than 1600 sq. centimeters. Compressive strength: The average compressive strength-in pounds per square inch of three mortar cubes composed of one part cement and 2.75 parts fine testing sand, by weight, shall be equal to or greater than the following:

Jan 1, 1941

The pressure of the gas in equilibrium with sphalerite has been determined in the temperature range of 680' to 825°C, using the Knudsen orifice method. A comparison of these experimental pressures with those calculated from thermal data and from other equilibrium measurements shows that the vapor above sphalerite is predominantly dissociated ZnS. Equations have been given for correctly calculating dissociation pressures using the Knudsen orifice method. It has been shown that the experimentally determined pressure is the same, whether the zinc sulphide is sphalerite or not, or a mixture of wurtzite and sphalerite. CONFLICTING points of view appear in the literature on the constitution of the vapor in equilibrium with solid zinc sulphide in the vicinity of 800°C. By comparing the dissociation pressure calculated from thermodynamic data and the vapor-pressure determination of ZnS by Veselovski,1 Lumsden2 has concluded that the vapor consists largely of dissociated ZnS. Sen Gupta,' however, concludes from his spectroscopic determinations that the vapor is largely ZnS molecules. In view of the fact that the thermodynamically calculated' dissociation pressure is higher than that experimentally measured by Veselovski, it seemed in order to repeat Veselovski's measurements. Experimental Procedure The method used for the determination of the pressures in this papel- is the Knudsen effusion cell. The apparatus and procedure were described in a previous paper- from this laboratory on the determination of the vapor pressure of silver. The only difference is that the Knudsen cell in this work is made from platinum and there is no external cover around the cell. The cell is an ordinary platinum crucible of 2.2 cm top diameter with a capsule cover. It was thought that platinum might stand up at these temperatures to the solid and gaseous ZnS, since it was found that the weight of the platinum cell itself did not change appreciably on heating ZnS in it at the working temperatures. To insure that reaction of the zinc sulphide with the cell was not giving' a false value, a stabilized zirconia cell was employed for check runs. Fig. 1 shows the comparison, which is satisfactory. Veselovski previously had measured the vapor pressure of ZnS using a silica Knudsen effusion cell. On repeating his experiment in this laboratory, it was found that ZnS at-tacked the silica cell, giving it a marked frosty appearance. This led to the belief that Veselovski's result:; may be in error. Also, he was operating at pressures above the range ordinarily considered safe for the Knudsen method. The effusion rate was measured by weighing the cell before and after each run. The weight loss during heating to temperature and cooling down was measured and subtracted from the weight loss during the actual run. The zinc sulphide used in this investigation was from two sources: Fisher cp grade, and a sample of pure sphalerite supplied by Mr. E. A. Anderson of the New Jersey Zinc Co. Before and after the series of runs with Fisher ZnS, X-ray analysis showed that both wurtzite and sphalerite were present. However, the ratio of sphalerite to wurtzite increased. All measurements were made below the transition temperature which has been reported" to be 1020°C. The data obtained in this investigation are tabulated in Table I. The pressure was calculated by the usual Knudsen formula" on the assumption that ZnS molecules were effusing. From these data, using pure sphalerite in the platinum Knudsen cell, the vapor pressure of ZnS, in mm of Hg, as a function of temperature is given by the solid line in Fig. 1. The best straight line, as determined by the method of least squares, is given by 14405 logpzns =-14405/T +11.032. A comparison of these results with Veselovski's shows that his results are about 50 pct lower. Discussion The vapor in equilibrium with solid zinc sulphide in the temperature range of this study will consist of Zn, S2, and ZnS mol, since other species of zinc and sulphur' are relatively unstable. The question to be settled is whether or not ZnS is largely dissociated. The derivation8 which follows gives the method of calculating the pressure of zinc and sulphur over solid ZnS, assuming complete dissociation, from Knudsen cell data. The free energy of the reaction 2 ZnS(solid) ? 2 Zn(gas) + S2(gas) is given by ?F?° = -RT In K = —RT In p12p2 where p1 is the zinc pressure and p is the sulphur pressure. If dissociation occurs in a closed system,

Jan 1, 1955

By F. A. Forward, A. H. Ross, J. Halpern

IN treating carnotite ores by carbonate leaching it is often necessary to subject the ore to a prior roast. Among the advantages that may result from roasting are: 1) improvement in settling and filtering characteristics of the ore; 2) destruction of carbonaceous impurities that would otherwise contaminate the leach solution and consume carbonate reagent; and 3) substantial increase in extraction of vanadium, which is usually low when the un-roasted ore is leached. Recovery of vanadium may be an important factor in processing carnotite ores, and the fact that vanadium is usually recovered in good yield by acid leaching has often been an economic advantage over results obtained with carbonate leaching. It has been found that response of an ore to carbonate leaching, following roasting, depends primarily on its composition and on the conditions of the prior roast treatment, particularly with regard to the roasting temperature, the atmosphere maintained above the charge during roasting and cooling, and the presence of certain reagents. These factors, especially the last, virtually determine the extractions of uranium and vanadium attainable by carbonate leaching, as well as the reagent consumption on leaching, whereas adjustment of the leaching conditions has only a limited influence. These variables were examined systematically to elucidate roasting and carbonate leaching of carnotite ores and to establish a treatment yielding high recoveries of both uranium and vanadium. Tests were conducted on samples of four different Colorado Plateau ores, described in Table I. Roasting Procedure: The roasting tests were conducted in an electrically heated furnace in which the temperature was controlled to +15oC. The ore (usually a 500-g charge) was ground to —100 mesh, mixed with the desired reagent, and spread in a silica tray to a bed depth of about 1/2 in. During the roast the charge was rabbled at intervals of 15 min. To provide free access of air the furnace was usually ventilated; however, in a few experiments a controlled atmosphere such as carbon dioxide or hydrogen was maintained above the charge. Leaching Procedure: The pulp, generally comprising 500 g of ore and a 2000-cc solution containing 80 g per liter of Na2CO3 and 20 g per liter of NaHCO3, was agitated in an autoclave at 120°C, under 30 psi O2. Samples were withdrawn periodically, and the solution and ore residue were analyzed. Under the conditions employed, maximum attainable extractions were generally achieved in 4

Jan 1, 1958

Commonwealth of Pennsylvania, Department of Internal Affairs, Topographic and Geologic Survey, Harrisburg, Pa George H. Ashley, State Geologist All available printed Bulletins may be obtained through the Commonwealth of Pennsylvania, Division of Documents. Where a price has been set on the publication, checks should be made out to the Commonwealth of Pennsylvania. A Syllabus of Pennsylvania Geology and Mineral Resources has recently been published by the Survey as Bulletin G1, and consists of a general survey of the state An appendix of this bulletin also gives a complete bibliography of Survey publications It will be sent upon request. The State Geological Survey has gone through several phases and reorganizations, publications of three of them being numerous and important A fire, in 1927, destroyed supplies of many of the papers of the present Survey. Available bulletins of the present Survey are: Bulletin M6: Bituminous coal fields of Pennsylvania, in 4 parts available as follows: pt 1, General description of coal, 50 cents; 3, Bituminous coal resources, 25 cents, 4, Analyses of bituminous coals, 25 cents Bulletin M10, Fire clays of Pennsylvania; Mll, Molding sands of Pennsylvania; M12, Anthracite culm and silt (1928); M13, Feldspar in Pennsylvania; M14, Magnetite deposits of French Creek (1931); County Report C1, Adams County, pt 1, Geology and geography (in press), County Report C2, Geology and mineral resources of Greene County (in press); Topographic and Geologic Atlas A5, New Castle Quadrangle (1929), $1; Topographic and Geologic Atlas A27, Pittsburgh Quadrangle (1929), $1; Topographic and Geologic Atlas A168, Lancaster Quadrangle (1930), $1. The present Survey has also issued a number of short mimeo¬graphed bulletins any of which will be sent upon request Briefly they are: No. 15, Mineral resources of Pennsylvania; 20, Iron ores; 28, Magnesite; 40, White clay deposits, Monroe County; 45, White clay deposits, central Pennsylvania; 47, Manganese, Eastern Pennsylvania, 48, Pyrite from bituminous coal mines; 59, Bog

Jan 1, 1933

By Willard C. Lacy

Rather than attempting to present a summary of the many and highly varied papers that have been presented at this symposium on sampling and grade control, I will attempt to extract the general philosophy of analysis and approach, and attempt to identify the trend of future developments. First, the term "sampling" is used with its broadest connotations. A sample consists of a representative portion of a larger mass, and must represent the mass not only in the grade of contained metals or minerals, but also in all other respects in terms of mineralogy and mineral quality (1, 5), deleterious materials, recoverability of economic components, physical behavior, geophysical response (I), and even archaeological and environmental aspects (7, 11). The sample must be taken from a locality and in such a manner and quantity that it is representative of the larger rock mass. This calls for complete and accurate geological control and an understanding of the nature and distribution of the contained chemical and physical elements and a record of the effectiveness of the different sampling methods. Second, value of a given mass of ore material is based upon its profitability - the difference between recoverable value and costs to achieve recovery, beneficiation and sale. There is a strong movement in mining geology control toward more complete analysis in determining cutoff grades and in grade control, as illustrated by the kriging of metallurgical recovery factors as well as grade at the Mercur Mine (8). To achieve a "profit- ability factor" as a guide for economic mining practice requires further integration of: 1) the value of contained metal or mineral, 2) percentage recovery of values, 3) dilution of ore with waste rock, 4) addition to, or loss of value as a consequence of by-product materials or deleterious components, 5) cost of producing a saleable product plus mini- mum profit to justify the effort (cutoff), and 6) cost of land restoration (7, 11). All these parameters vary with the rock type, rock structure, mineralogy, depth, geometry, mining and metallurgical methods, but they must be sampled and analyzed if sampling and grade control are to reflect profitability. A wide variety of deposits has been presented at this symposium; each deposit with its own problems and special solutions. Deposits containing high unit-value components, e.g. precious metals and diamonds, present special problems in the obtaining of accurate samples and generally require statistical analysis control methods or may disregard or modify occasional high or occasional low values, based upon experience (12 ) Grade control may be accurate for the long term but may vary for the short term. Bulk sampling is always essential. Deposits containing metals or minerals with low unit value are very sensitive to transport costs, and they are often very sensitive to small amounts of deleterious components or differences in physical or chemical behavior. Problems of sampling and grade control change with the genetic type of deposit, with the stage of deposit development and with the size of the information base. Precious metal epithermal deposits (2, 6, 8), because of rapid vertical zonation and erratic lateral distribution of values, have always been difficult to evaluate and maintain grade control and ore reserves. On the other hand, evaluation and grade control are relatively easy in bulk-low- grade deposits (4, 13). However, these deposits generally have a low margin of profit and are sensitive to mining and beneficiaton costs, price fluctuations and political costs. Industrial mineral deposits (5) often must be evaluated on the basis of their behavior, rather than by chemical analysis. Environmental impact generally increases with the scale of the operation, but certain elements or minerals have especially high impact effects (7, 11). In the exploration phase there is no production control of sampling procedures and careful geological observations are particularly essential. The greatest number of problems is related to the oxidized outcrop where the chemical environment of the ore body has changed and the contained values may have been enriched, depleted or values left unchanged (2, 6). Present evidence suggests that gold values may be very mobile under certain conditions (2, 6) and stable under others. Everything must be sampled in detail. Principal values and by-product or deleterious elements may vary dependent upon their position within the soil profile. Such factors as geomorphic position, erosion rate, vegetation, climate, etc., may affect the interpretation (1, 3). During the development phase it is equally easy to overtest, to have "paralysis by analysis," as to undertest (3, 6). Bulk samplng and testing are

Jan 1, 1985

By Philip R. Strand, O. F. Stewart

Vermiculite is the name generally applied to the group of hydrated ferromagnesian aluminum silicates that are characterized by the ability to expand when heated. This process, called exfoliation, results in a lightweight product of commercial value. Most uses of vermiculite are for the expanded material. The chief markets are in construction, agriculture, and horticulture, with lesser uses in general industry. Composition and Properties Vermiculite, in its natural state, has the characteristic micaceous habit, a perfect basal cleavage which causes splitting into thin laminae that are soft, pliable, and inelastic. The structure of vermiculite is basically that of a talc. The prominent monoclinic crystal faces are often marked by lines at 60° and 120°. Hardness varies from 1.5 to 2 or more; specific gravities are between 2.1 and 2.8; color varies from almost clear to amber, bronze, brown, green, or black. Vermiculite feels like talc, especially when wet. Although much research has been performed on the chemical and structural composition, there is not yet complete agreement on the exact formula. This is to be expected when different workers have examined the many different varieties. Vermiculite is not considered to be a single mineral species but a family of related minerals. The structural formula for a trioctahedral vermiculite may be written: (H20)-(Mg,Ca,K)-(A12,Fe,Mg)-(Si,A1,Fe)4O10 (OH) 2 Hydrobiotite also occurs with vermiculite and is usually considered a vermiculite for commercial uses. When heated quickly to elevated temperatures, vermiculite expands by exfoliating at right angles to the cleavage into wormlike particles. The name vermiculite is derived from the Latin vermiculare, to breed worms. This characteristic of expansion is the result of the mechanical separation of the layers by the rapid conversion of contained water to steam. The decrease in bulk density of commercial grades is usually approximately 10 times, from 50 to 5 lb/ft3, but varies depending on the quality, size, and furnace efficiency. Individual flakes may expand up to 30 times. Vermiculite may also be expanded by soaking in chemicals such as hydrogen peroxide, weak acids, and other electrolytes. Color change during expansion is dependent upon the type of vermiculite and furnace conditions. Heating in an oxidizing atmosphere produces a dull gray or tan color, whereas a reducing atmosphere can produce a bronze or gold color. The expansion of the vermiculite crystal results in large pores being formed between the platelets. Thus, exfoliation makes available a large increase in void volume which is important in the application of vermiculite as a chemical carrier. Mathieson and Walker (1954) state vermiculite must be regarded as a true clay mineral. The characteristic properties of the mineral, such as high cation exchange capacity, organic complexing ability, and variable interlamellar distance are very similar to those of montmorillonite. The cation exchange capacity of vermiculite is one of the highest of all the clay minerals. The interplatelet space is accessible to penetration by some electrostatically neutral molecules. Water and glycerine are two common substances whose molecules may be so imbibed. In the natural state and under normal atmospheric conditions, water occupies the spaces between the silicate layers. The crystal d-spacing is near 14.2A. By differential thermal analysis, it has been determined that the water

Jan 1, 1983

By Thomas Fraser, H. F. Yancey

The jig may be termed the standard coal-washing machine. Although exact figures on the relative tonnages of coal treated by the various coal-cleaning processes are not available, it is known that a large part of all the coal given special cleaning treatment is handled by jigs. This wide use of the jig warrants a special study of its performance. In this paper are presented the results of an analysis of work done by a three-compartment jig in treating bituminous coal from the Thick Freeport bed in Allegheny County, Pa. Certain systems of washing coal require sizing before washing, while others treat unsized coal. The prevailing practice, however, for the preparation of coking coal from the more easily washed coal is to treat an unsized or natural feed of about — 1 in. size and smaller. Under such conditions, the minimum size of particle that is beneficiated has long been and is still a matter of discussion and conjecture. In a careful laboratory investigation J?ngstl concluded that the practical limit of settling of fine coal is reached at from 0.02 to 0.008 in. On the other hand, Hancock2 has found that single-compartment jigs, used in preparing coal for the domestic market, usually allow most of the heavy (refuse) particles smaller than 1/4 in. to go into the washed coal. He also states that it is impossible with the ordinary coal jig to make any noticeable improvement in the quality of the coal finer than 20 mesh (about 1/30 in.). Because of its importance in coal cleaning, the work described in this paper was undertaken for the purpose of obtaining information on this subject, as a part of the general coal beneficiation program being carried out by the Bureau of Mines.

Jan 1, 1925

By P. Dosmond

In ore mining technology, the loading and hauling operations are undoubtedly the area where the most spectacular progress has been made during the last twenty years. In the iron ore mines of Eastern France production of 62 million tons for 1962 and 19863 constitute a worthwhile experience worldwide. These deposits which are located in Lorraine are of sedimentary origin and they contain many mineralized horizons with workable seams varying in depth from 330 to 830 ft.

Jan 1, 1970

By S. W. Parr, T. E. Layng

The low-temperature carbonization of coal involves the carrying out of the coking process under conditions wherein neither the coal mass nor any of the passageways through which the volatile products pass are heated above 700" or 800" C. For convenience in this discussion, the single number 750" will be used to designate the maximum range. This temperature is not selected arbitrarily; it is the result of certain natural conditions that are inherent in the substances involved. Two of these conditions arc sufficiently pronounced to suggest a line of demarcation at this point as follows: (1) Below 750°, all the heavy hydrocarbons are expelled, which means that, at these lower temperatures, the illuminants, the gases of high calorific value, and the condensible oils are discharged; above 750") there are given off the lean, non-illuminating gases consisting for the most part of hydrogen and marsh gas and having no condensible constituents present. (2) Below 750") there is substantially no secondary decomposition; above 750") the volatile products are readily decomposed, forming tars, naphthalene, free carbon, etc. It is not intended to maintain that no secondary decompositions occur below 750". Many recent studies have demonstrated the practicability, especially in the presence of catalytic substances, of cracking certain of the hydrocarbon compounds; but at these lower temperatures the step is a moderate one, as, for- example, from xylene to toluene or from toluene to an anthracene. These changes are moderate in amount. Not only do the reactions proceed slowly but they are subjected to the decomposing . conditions for only a short time. This is evident when it is recalled that at these initial temperatures the decomposition of the coal is very rapid and, if anywhere near a neutral pressure is maintained, the movement of the evolved gases is lively and reduces, correspondingly, the time for the retention of the gases in the passageways where these milder decomposing conditions exist. It is acknowledged that the maintenance of these temperature conditions at the present time has only an ideal and not a practical status. Actual operations under these conditions, as a? industrial accomplish-

Jan 1, 1920

By H. C. Gatos, J. T. A. Pollock

It is well known that the superconducting critical current densities of many alloy superconductors may be increased by cold working and in some cases further enhanced by a short heat treatment. This latter enhancement has been attributed to the redistribution of dislocations into cell-like networks' and to the precipitation of second phase particles,2'3 which act as flux pinning centers. In a manner analogous to dislocation pinning in precipitation hardening alloys,4 it is expected that here also a critical distribution of the pinning centers should result in maximum pinning effect. Concentration inhomogeneities exist in most or all commercial alloys yet there have been only a few attempts made to determine their effect on critical current capacity in the absence of cold working. Sutton and Baker,5 and Kramer and Rhodes6 have found that the complex precipitation processes occurring during the aging of Ti-Nb alloys can result in critical current density enhancement. Livingston7-10 has clearly shown, for lead and indium based alloys, that the distribution of precipitated second phase particles is of critical importance in determining magnetization characteristics. However, these '(soft" alloys age at room temperature and the time involved in specimen preparation prevents metallographic examination in the state in which the superconducting measurements are made. Thus results with such alloys are expected to be biased towards larger precipitates and interpar-ticle spacing. The present study of Ta-Zr alloys was undertaken to examine the influence of second phase precipitation, as controlled by heat treatment, on the critical current capacity of well annealed polycrystalline material. A study of the published phase diagram11 indicated that annealing supersaturated samples containing up to 9 at. pct Zr at suitable temperatures would result in the precipitation of a zirconium-rich second phase. It was MATERIALS AND PROCEDURE The alloys were prepared from spectrochemically pure tantalum and zirconium. Analysis was carried out by the supplier. Major impurities in the tantalum were: 12 pprn of 02, 17 pprn of N2, 19 pprn of C, and less than 10 ppm each of Mo, Nb, Al, Cr, Ni, Si, Ti. The crystal bar zirconium was pure except for the following concentrations: 15 pprn of 02, 17 ppm of C, 23 ppm of Fe, 11 ppm of Cu, and less than 10 pprn each of Al, Ca, N2, Ti, and Sn. Samples were prepared in the form of 8 to 10 g but-tons by arc melting using a nonconsumable electrode on a water-cooled copper hearth in a high purity ar-gon atmosphere. Each button was inverted and re-melted three times to ensure an even distribution of the component elements. The samples were then homogenized at temperatures close to their melting points for 3 days in a vacuum furnace maintained at 5 x 10-7 mm Hg. After this treatment the buttons were cold rolled to sheets approximately 0.020 in. thick from which specimens were cut, 0.040 in, wide and 1 in. long suitable for critical current density (J,) and critical temperature (T,) measurements. These strips were then recrystallized and further grain growth was allowed by an additional vacuum heat treatment at 1800°C for 60 hr. Some second phase precipitation occurred during cooling of the furnace and a solution treatment was necessary to produce single phase supersaturated samples. This treatment was successfully carried out by sealing the samples together with some zirconium chips in quartz tubes under a vacuum of 5 x 10-7 mm Hg, heating at 1000°C for 5 hr and then quenching into water or liquid nitrogen. The samples were then heat treated at either 350" or 550°C and quenched into water or liquid nitrogen. All samples which were heat treated at 350°C were quenched in both cases by cracking the capsules in liquid nitrogen. The samples treated at 550°C were quenched by dropping the capsules into water. Analysis for oxygen in randomly selected samples indicated that the oxygen content was in the range of 175 to 225 ppm. Values of Tc were determined by employing a self-inductance technique. Jc measurements were made at 4.2oK by increasing the direct current through the wire in a perpendicularly applied field until a voltage of 1 pv was detected with a null meter. The risk of resistive heating at the soldered joints during these latter measurements was reduced by first plating the ends of the wires with indium and then soldering to the copper current leads using tin. Metallographic examinations were performed after mechanical polishing of the same samples and etching in a 4H20:3HN03 (conc):lHF(conc) solution.

Jan 1, 1970

By R. D. Satterley

THE Sherwood mine is an iron-ore mine owned and operated by the Inland Steel Co. in the Iron River district of the Menominee Range in Michigan. The property consists of an 80-acre tract in the village of Mineral Hills, about a mile north of the city of Iron River. The ore mined is a reddish brown, nonbessemer hematite and limonite. The ore bodies lie in a large syncline having an east-west strike and are associated with a rock structure that is folded in a complex manner. The two 40-acre tracts run north and south and the shaft is at about the center of the north forty. The shaft is 1286 ft. deep, with the principal operating level at a depth of 1000 ft. from surface. A new level at a depth of 1200 ft. is in process of development. The ore body at the 1000 ft. level is east and west at approximately the line between the north and south forties and at this elevation is from 50 to 250 ft. wide and extends above the level at varying heights of from 50 to 250 feet. The 1000-ft. haulage drift extends south from the shaft, through the east-west ore body to the slates on the south side of the body. The ore body was intercepted in the drifting at about 800 ft. from the shaft and was about 200 ft. wide at that point. A crosscut was extended to the east and west property lines along the south boundary of the ore body. This east-west ore body is laid out in areas 80 ft. wide for stoping with pillars 40 ft. wide. With this plan a stope is approximately 80 by 200 ft. and has a pillar on both sides that is approximately 40 by 200 ft. The ore has a height of from 50 to 250 ft. above the elevation of the level, therefore the height of the stopes varies in accordance with the height of the ore body at each location. From the crosscut, two main raises, each 6 by 7 ft., are put up into each 80 ft. stope area (Fig. I). One of the main raises serves as a manway and the other as an ore raise, through which the ore is transferred from the stope to the tramming equipment. The original layout was for scraping of the ore from under the stopes into the ore raises. The ore raise is put halfway (east and west) along the 80-ft. stope area and usually is in iron formation. The manway and ore raise are on approximately 15-ft- centers. These main raises are put up to a height of from 22 to 30 ft. above the level floor, then from the top of the ore raise a scraper drift (8 X 8 ft.) is driven across the width of the ore body, which is in the neighborhood of 200 ft. From the scraper drift, mill raises are put up at 12 ½ -ft. centers, staggered on opposite sides of the drift, and these are extended 23 ft. above the elevation of the scraper-drift floor. The sublevel connecting the top of these mill raises is called a mill "sub." Other sublevels are developed parallel to the scraper and mill subs at 25-ft. vertical intervals above the mill sub. The mill sub and the other subs above the mill sub are known as sublevels or dog drifts, and are 4 ft. wide and 6 ft. high. These dog drifts

Jan 1, 1945

By Thomas C. Wilder

The concentration and activity of oxygen in liquid copper has been measured by the reversible galvanic cell: at 1100°and 1200°C for the concentration range of oxygen of 3 to 5600 ppm Measurements for approximately thirly alloys at each temperature indicate that the activity of oxygen roughly follows Henry's law, the activity coefficients of oxygen being 0.12 and 0.205 at 1100° and 1200°C, respectively. These results compare javorably with the results of seven other electrochemical and gaseous equilibrium studies on this system, the free energy of the solution reaction 1/2 O2 = O (7 at. pct in Cu, liq) being -18.10 * 0.30 kcal per gfw at 1200°C. Estimates of the standard molar free energy of formation of CupO(S) are also made. Praclicul consideration of this cell as an oxygen sensor for liquid metals is discussed briefly. SOLID mixed-oxide electrolytes which have a defect structure in the anion sublattice have recently become a very popular means by which the chemical potential of oxygen is measured at high temperatures. Since the pioneering work of Kiukkola and Wagner1 such electrolytes have been used in numerous cell studies for determining the thermodynamic properties of metallic oxides and oxy compounds, the partial pressure of oxygen in vacuum systems and other atmospheres, and the properties of oxygen dissolved in metals and alloys. The use of a solid electrolyte in a cell for a fast reliable method of oxygen determination in liquid metals would be a very important tool to the process metallurgist. In copper, for example, the oxygen content of the molten metal in the cathode furnace is a prime factor in determining the "set" or solidification profile of cast wire bars. In this investigation the activity of oxygen in liquid copper was measured by an electrochemical cell employing the electrolyte ZrO2 + 7 1/2 wt pct CaO. The cell used at 1100° and 1200°C was

Jan 1, 1967

By D. A. Elgillani, S. Atak, D. A. Rice, M. C. Fuerstenau, R. B. Bhappu

Quartz and feldspar cannot be floated with sulfonate at any pH; spodumene floats over a narrow acid pH range, while beryl responds moderately over a broad pH range. After wet-grinding in a steel mill, beryl, quartz, and spodumene float well with sulfonate below about pH 7, whereas the improvement in the response of feldspar is not so marked. A mechanism by which iron can be adsorbed on these minerals is presented. Also, the responses of leached, natural, and wet-ground beryl to amine, sulfonate, and oleate flotation are shown and related to the measured zero-points-of-charge of these materials. Earlier work with leached beryl showed that good flotation could be obtained with alkyl aryl sulfonate over a rather wide pH range using a Fagergren flotation cell.' When a similar response was observed with leached quartz, it was decided that unintentional activation was being obtained from the metallic components of the Fagergren cell. To obviate this difficulty, a microflotation cell was designed, and an experimental technique was devised. These have been described elsewhere. Experiments conducted with the small cell showed that leached quartz could not be floated at any pH with any sulfonate addition,3 which is in agreement with the observations of Kraeber and Boppel.4 Similarly, it was also found that leached beryl responded to sulfonate flotation only over a narrow pH range rather than the broad range reported earlier.1 This early work,1 however, revealed the important effect that wet-grinding in a steel mill has on the flotation response of certain silicates. That is, it was found that quartz and especially beryl floated well over an unusually wide pH range after wet-grinding in a steel mill. Microcline, however, floated poorly below pH 4, even though wet-ground under the same conditions. The work of Eigeles6 on adsorption of oleic acid on leached quartz and iron-contaminated quartz at constant pH is in agreement with these flotation data. Other research has shown that ferric iron, added as a salt to the system, functions as an activator in the narrow pH range in which Fe +++ iron hydrolyzes to its hydroxy complexes.3,5 These phenomena indicate that iron functions differently in flotation systems depending on its method of introduction. The object of this paper is to determine the mechanism by which iron is adsorbed on certain minerals, the mechanism of collector adsorption after iron abstraction, and the role that Fe++ and Fe+++ assume in the selective separation of these minerals. EXPERIMENTAL MATERIALS AND METHODS Sodium alkyl aryl sulfonate, mol wt 450,7 pure potassium oleate, and pure dodecylamine were used as collectors. All other chemicals were reagent grade in quality, i.e., n-amyl alcohol as frother; HC1, H2SO4, and KOH for pH adjustment; and ferric chloride as activator. Conductivity water, made by passing distilled water through an ion exchange column, was used in the experimental work. All minerals used in the investigation were hand-picked specimens. Sample Preparation: Each of the minerals was crushed through 8 mesh, and the product was divided into two groups, one to be ground dry and the other wet. Dry grinding was accomplished with an alumina mortar and pestle. The product was dry-screened to 48 x 150 mesh, cleaned magnetically, deslimed in conductivity water, and dried. Preparation of the samples by wet-grinding involved grinding a 200-g charge of the mineral (-8 mesh) at 60% solids with natural water in a mild steel rod mill for four minutes. This charge was then wet-screened immediately with natural water to 48 x 150 mesh, dried, and cleaned magnetically. Some experiments were also conducted with leached beryl and quartz. These products were prepared by leaching the sized sample (48 x 150 mesh) with concentrated HC1 with a percolation technique until no iron could be detected in the leach liquor. Following this step, the sample was rinsed with conductivity

Jan 1, 1967

By Mark Sheppard

DURING the winter of 1937, the writer visited a West Virginia stone, quarry at which the overburden is stripped hydraulically. The quarry is in a bed of limestone, about 200 ft. thick, which outcrops on a hillside above a small river; the quarry face is 175 ft. high. The top of the stone is cut in all directions by erosion channels, which are filled with a tough, red clay. Some of the larger channels are 30 ft. deep and 8 ft. wide, The thickness of the clay overburden ranges from a few inches to 10 ft., the average being estimated as less than 3 feet. The topography of the district is rugged, and the hillside above the quarry face is not suited to the use of the usual mechanical methods of moving dirt. Until hydraulicking was started 10 years ago, the over-burden was dug by hand labor, dumped to the quarry floor and trans-ported to the dump in quarry cars, Cleaning deep channels was a slow and expensive process. When hydraulicking was started, the top of the quarry face was over 200 ft. above the level of the river. The pipe line was carried in a semi-circle around the top of the quarry about 200 ft. from the face and about 50 ft, above it. As there is no adequate supply of water above the quarry, the water must be pumped from the river: As quarrying pro-gressed the pipe line was moved until the discharge points are now over 300 ft. above the river. Water is pumped from the river by a four-stage centrifugal pump having a 5-in. intake and a 4-in. discharge. It is direct-connected to a 200-hp,, 2300-volt, 1760-r.p.m. induction motor. At 1760 r.p.m. the rated capacity of the pump is 750 gal. per minute under a head of 600 ft. The pump is installed in a house of frame construction on the river bank, about 10 ft, above normal water level, which also houses the pump for the mill-water supply. Above the pumps are hoisting hays into which the motors are raised whenever the river reaches flood stage.

Jan 1, 1938

By Walter Rose

It is the purpose of this note to call attention to the circumstance complicating the attainment of uniform gas-liquid distributions in multi-phase flow systems, and especially in those of the so-called Hassler type.' Although the complication arises directly as a consequence of gas compressibility and the dependence of fluid distributions on interfacial curvature phenomena, little reference to the problem has yet appeared in the literature.' In fact, it is only the paper of Geffen et al.3 which gives any experimental evidence that a problem exists. On the other hand, previous authorsl,4,5 have made the tacit, albeit fallaceous assumption that uniform gas-liquid distributions automatically are established in linear flow systems by the expedient of maintaining the pressure gradients equal in the flowing gas and liquid phases. Such a device, it can be shown, will succeed only if the immiscible fluids are both equally compressible or both essentially non-compressible. Another necessary condition required before uniform fluid-fluid distributions can be achieved is that the porous matrix is isotropic. Uniform fluid saturation distribution conditions obtain in isotropic media only when the curvature of each interface of contact between wetting and non-wetting fluids is everywhere the same throughout the interspaces. This is a necessary condition which follows from the consideration that variation in interfacial curvature gives rise to finite capillary pressure gradients and therefore to variation in the saturation distribution. That this is not a sufficient condition follows from the consideration that hysteric possibilities allow for different saturations even though the capillary pressure gradient is zero. In any event, it will- be recalled that it is the intent in the Hassler scheme of relative permeability determination' to obtain initially (by the capillary pressure drainage or imhibition process) uniform conditions of fluid distribution in the core sample, and then to maintain this uniformity during mixture flow so that the resultant fluid mobilities which are calculated will refer to steady-state transfer under fixed conditions of uniform saturation. If the flowing fluids are incompressible, it is recognized that employment of the same value of the pressure gradient in each fluid will result in maintenance of the initially obtained condition of zero capillary pressure gradient. On the other hand, the consequence of gas compressibility is that the gradient in the gas pressure varies from point to point in the flowing stream. and since the pressure gradient in an incompressible liquid is constant it is thus impossible to maintain zero capillary pressure gradient during the simultaneous flow of gas-liquid mixtures. Therefore, one would expect to observe a variation in the gas-liquid saturation in the directions of the capillary pressure gradients, depending in magnitude on the particular way saturation changed with capillary pressure. In order to formulate the manner gas compressibility effects prevent the attainment of uniform gas-liquid distributions during mixture flow, consider a linear flow system of unit length. As noted, it is necessary to assume an isotropic medium (viz. one where permeability is independent of position and direction). One then can arbitrarily set the pressure

Jan 1, 1951