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By William P. Blake

This camp is just now startled by the alarming suggestion that unless there shall speedily be an appreciation in the value of silver, and a decrease in the cost of salt, it will be prudent business policy to close the Alice mines and mills until the conditions of production are more favorable. Such an event would be a great blow to the prosperity of Butte. It would throw some hundreds of men out of employment, and it might lead to the adoption of the same policy by other companies, which, like the Alice, have not been able to pay dividends to their shareholders for a long time. The question arises, and becomes emphatic: If, under the present adverse conditions, the mines and mills cannot be made to pay a proper profit, why should they continue to be worked ? The owners mentally ask: Shall we go on exhausting our mines, pouring out a stream of silver for the maintenance of the community, and the support of the rail ways, or shall we husband our reserves, withdraw the miners from our slopes, and wait for more favorable conditions ? This is the impending question, and all concerned dread to meet and answer it. Let us briefly review the situation and forecast, if possible, what the solution of' the problem will be. It must be understood that the discussion does not include, or refer to, the immense copper-producing interests of the camp, which may or may not be upon a very much more satisfactory basis. It will be confined chiefly to the operations of the Alice company, sta tistics of which are at hand, without involving the other silver com panies in the conclusions beyond those inevitable from well-known facts, patent to all. There are at present in Butte and in its vicinity in the principal mills an aggregate of 290 stamps dropping upon silver-ores: Alice,...........80 stamps. Moulton,. .........40 " Lexington,...50 " Silver Bow,.........30 " Blue Bird,..........90 " The Alice mills at present, with their 80 stamps, devour, in round numbers, from 100 to 120 tons a day. The actual amount

Jan 1, 1888

By Leroy U. Gardner

EVERY reader of this paper is well aware of the fact that the prolonged inhalation of large amounts of free silica dust results in fibrosis of the lungs, and that other inorganic dusts, except those of asbestos and the mineral poisons, are thought to be relatively harmless. But probably few have had opportunity to visualize the effects of silica, asbestos and other minerals upon living tissues. The ideal demonstration would be based upon sections of lungs from human beings that had inhaled com- parable concentrations of different dusts for a unit period of time. Such material is generally difficult to obtain and in some instances is of questionable value because the subject has been exposed to more than one kind of dust. A demonstration could be prepared from the lungs of animals. However, so many months or even years are required to complete the exposure to each kind of dust that enough material has not been accumulated. For a general survey of the action of different types of dust, I am therefore compelled to use the tissues of animals that have been injected with suspensions of dust particles. The injection method offers certain advantages. The quantity, the particle size and the state of dispersion can be controlled with sufficient accuracy so that the tissue responses to different minerals can be accurately compared. The method insures contact between the tissues and the mineral particles; protective mechanisms, which may prevent the inhalation of certain types of foreign body, never come into play. If the pathway of injection be one that minimizes nonspecific injury, the tissue responses may simulate those produced by inhalation but they develop somewhat more rapidly. It should always be remembered that injection methods merely demonstrate the capacity of various kinds of minerals to provoke reaction in the tissues; whether the same substances are inhaled from atmospheric suspension and accumulate in the lungs in dangerous quantities can be demonstrated only by inhalation. The reactions that I shall discuss first took place not in the lungs but in the liver. For the purpose in hand this is quite immaterial for we are merely seeking to discover whether these minerals are capable of produc-

Jan 1, 1938

By Walter L. Maxson, Fred C. Bond

THIS paper is a continuation of two earlier papers, l, 2 and presents new data on the grindability of various ores and other materials-the results of several years of intermittent research work on the nature of grinding and interpretation of test results. It is divided into five parts: 1. The tabulated results of several hundred grindability tests on a wide variety of ores. 2. A method of correcting screen analyses and calculating from them the surface and other properties. 3. A description of a new method of measuring the "surface energy" of an ore by pendulum impact tests. 4. Test results. 5. A general discussion of the test results and their application to grinding problems. 1. GRINDABILITY OF VARIOUS ORES This part is a continuation of the paper listed as reference No. 1. The method of testing is described in detail in that paper. The sample to be tested is crushed to all through 6 mesh in rolls set at x6-in. opening in closed circuit with a screen. Of this sample 700 C.C. is ground dry in a laboratory ball mill under constant conditions of mill speed and ball charge in closed circuit with a screen of the desired size at 250 per cent circulating load. The mill is cylindrical, 12 by 12 in., and rotates at 70 r.p.m. It contains a charge of 285 iron balls weighing 20,125 grams and ranging from 1 ½ to ¾ in. in diameter. Each test is continued for a sufficient number of periods to establish equilibrium at 250 per cent circulating load, or until the composition of the circulating load is constant for successive periods. Tests are conducted at the mesh size to which the ore is to be ground in practice, and the results are expressed as the net production in grams of screen undersize per revolution of the mill. This is designated as the

Jan 1, 1938

By D. Webster

It has been found that cold work and annealing treatments greatly retard subsequent grain growth and re-crystallization in thoria-dispersed nickel and NiCr. It is suggested that this is due to cavities which form at the particle-matrix interface durilzg cold work. Some of' the cavities formed are stable at elevated temperatures and prevent the mouement of large-angle grain boundaries. The movement of these boundaries at elevated temperatures is known to start recrystal-lization and when they are pinned by voids recrystal-lization is made more difficult. TD Nichro,me had a maximun resistance to recrystallization when the initial structure was coarse-grained and had a low dislocation density. A precipitation reaction occurred between 700° and 1200° F in cold-worked TD Nichrome. It has been shown by Grant and his coworkers1-3 that a large part of the strength of dispersion-hardened materials is a result of the cold-worked structure produced during fabrication. In a comparison of the strength of cold-worked copper with and without an oxide dispersion, Grant4 has demonstrated that there is little difference in strength at room temperature. The function of the dispersion is essentially to maintain the cold-worked structure at higher temperatures than is possible in the base material. The dispersion accomplishes this by making recrystallization more difficult. The mechanism by which this is achieved is discussed in this paper. The fabrication history of the material has been found to greatly influence recrystallization behavior and this aspect is discussed in some detail. I) MATERIAL Three different thoria-dispersed materials were examined. 1) DuPont 2 pet thoria-dispersed nickel (TDNi) in 0.050-in. seeet. Thoria particle size is in the range 100 to 2500A. 2) DuPont 2 pct thoria-dispersed nichrome (20 pet Cr, 80 pct Ni) (TDNiC) in 0.125-, 0.060-, and 0.040-in. sheet. In the as-received condition these sheets have

Jan 1, 1969

By M. C. Chang, J. Dasher

The Crucible mine in Pennsylvania, operating on Pittsburgh seam coal, is rated at 5000 tpd. The washing plant, built in 1943, is rated at about 400 tph, using hydroseparator boxes to wash the coarse coal (4 x 5/16 in.) and hydrotator units for the fine coal (5/16 x 0). It is generally acknowledged that gravity concentration equipment cannot be used efficiently to bene-ficiate coal finer than about 65 mesh. Before the present flotation circuit was installed, Crucible recovered its fine coal slurry, predominantly —24 mesh material, by passing it through a battery of five 14-in. and one hundred and ten 3-in. Hey1 & Patterson cyclones in parallel with a 75-ft Dorr thickener. About two thirds went to the cyclones and one third to the thickener. Underflows from cyclones and thickener were filtered, respectively, in one 12-disk Eimco and two 4-disk Oliver filters fitted with stainless steel cloth. The resulting cakes were added to the coarse metallurgical coal. The filtrates, overflow from the 3-in. cyclones, and thickener overflow were recycled to the coarse coal plant. The three major problems encountered in this system became acute when the mine switched from track-mounted machinery to the more economical off-track equipment. The quality and uniformity of product declined, too much slime accumulated in the circuit, and waste pond disposal became highly impracticable. Lowering of Quality and Uniformity: Since the filter cake was not beneficiated in any way, addition of this material, which contained about 18 pct ash. would certainly lower the quality of metallurgical coal. Difficulties of handling the high-ash, high-moisture cake were compounded by the lack of adequate blending facilities at the mine, making it impossible to maintain uniform feed for the coke ovens. For a two-month test period, to verify the source of trouble, all filter cake was excluded from the metallurgical coal. The excluded filter cake was stocked in the yard. Slime Buildup: The moisture retained in the coarse coal products and the filter cake provided daily the only means of slime bleed from the circuit. As this bleed was not enough to offset the incoming fines in the raw coal feed, there was excessive buildup of slimes. To operate at all, it was necessary to purge the system each weekend by pumping all the solids from the thickener to a waste pond. The slimes built up during the week progressively increased the apparent viscosity and density of the circulating plant water, so that hydroseparators and hydrotator units had to be constantly adjusted. Uniformity, yield, and quality of the metallurgical coal suffered in consequence, and the filter cake ash and moisture contents increased from Monday to Friday. Limited Waste Disposal: Location of the plant does not permit unlimited waste pond area. Maintenance of a waste pond for the purpose of emptying the thickener during weekends became a costly burden that could not be tolerated. HOW THE PROBLEMS WERE SOLVED The flowsheet (Fig. 1) chosen to overcome these difficulties is considered the one most economical to install and operate. In a preliminary change made before the flotation circuit was installed, the cyclones and thickener were first operated in series, the overflow from the 3-in. cyclones, after flocculation, being discharged into the thickener and the thick-

Jan 1, 1960

By B. Bernstrom

A 22-ft diam, 7-ft long, wet autogenous grinding mill was installed in the new Cretaceous plant of the Jones & Laughlin Steel Corp. to prepare crude iron ore for concentration in spirals and flotation cells. The paper describes the flowsheet and gives a description of the operating and metallurgical results that have been obtained. These results include capacity, horsepower, operating problems and a comparison of pilot plant results with those obtained in commercial operation. The Cretaceous plant at the Hill Annex mine of the Jones & Laughlin Steel Corp. was designed with a wet autogenous mill for grinding low grade iron ore so that it could be concentrated by spirals and flotation cells. The plant, on the west end of the Minnesota's Mesabi Range, went into operation in June 1961, and preliminary operating and metallurgical information describing the performance of this mill can now be given. The crude is lean ore from the Cretaceous iron foundation. This formation is a beach deposit of Cretaceous age that overlies the Precambrian direct shipping and concentrating ores. This rock in nature is not extremely hard but is quite resilient, resisting blasting in mining and size reduction in milling. It contains hematite as water-rounded particles and other fragments of the underlying iron formation and adjacent rocks cemented together by a high alumina-silica ferruginous matrix. It is interspersed with hard dense hematite agglomerates and also contains granite pebbles and boulders that were not cleaned off during removal of the surface. It has a 3.48 sp gr and in the stockpile the material consists of frag-

Jan 1, 1962

By W. D. Sylwestrowicz

In tests on polycrystalline copper and aluminum, the ratio of the yield stress to modulus of elasticity was found to be strongly dependent on tempemture. Also, it was shown that the change of the yield stress with temperature of a material at a given temperature and deformed to a given yield stress, but not necessarily to the same strain, does not depend on previous mechanical thermal history. Likewise, for a material at a given temperature and deformed to a given yield stress, the rate of work hardening is independent of the temperature of prior stressing. It has been shown by Los and Orowan1 and by Dorn, Goldberg, and Tietz2 that the temperature dependence of the yield stress-strain curves of crystalline materials reflects two effects: not only does the yield stress of a material in a given structural state depend on the temperature, but the structural state (and the corresponding strain hardening) after a given amount of plastic straining also depends on the temperature at which the pre-straining has been carried out. Thus, for instance, if a specimen of a ductile metal has been strained along the stress-strain curve 1 at room temperature to point A, Fig. 1, and then the straining continued in liquid nitrogen along curve NR, the yield stress A' immediately after lowering the temperature is lower than the yield stress A" of a specimen that has been strained by the same amount in liquid nitrogen (along curve 3). AA' represents the effect of the temperature change upon the yield stress of the specimen pre-strained at room temperature; A' A" is the excess strain hardening acquired if the pre-straining has been carried out in liquid nitrogen instead of at room temperature. The dependence of the yield stress (i.e., of the structural state acquired during preceding straining) upon the conditions (speed, temperature) of a pre-straining is the reason for the nonexistence of a so-called "mechanical equation of state" for plastic deformation. A 'mechanical equation of state" is a relationship of the form F(s,E, e, T) = 0, or s = f(E, E, T) [1] It was first suggested by Ludwik3 that plastic deformation might obey such an equation; this, however, was conclusively disproved by experiments of Kochendorfer.4 Additional disproofs were given in

Jan 1, 1959

By Benjamin LeRoy Miller, Joseph T. Singewald

For a number of years Russia, India and Brazil have outranked all other countries as producers of manganese ores. During the 5 years immediately preceding the European war, the average annual production of Russia was 844,000, of India 694,000 and of Brazil 200,000 long tons. Since the outbreak of the war there has been a considerable falling off in the Russian and Indian production, particularly in the former, whereas the production of Brazilian ores has been greatly increased, amounting in 1914 to over 250,000 tons and in 1915 to nearly 350,000, with conditions favorable for a still larger production in 1916. Manganese is one of the few industrially important metals that are not produced in the United States in quantity commensurate with our needs, so that we have been compelled to import annually about 300,000 tons, having a value of over $2,000,000, and these ores have been obtained from the three countries mentioned above. Their rank as contributors to our imports of manganese ores was India, Russia, Brazil; and in 1913, Brazil contributed only one-fifth. During the past 2 years Brazil has furnished rapidly increasing quantities to this country, and, with the falling off of imports from India and Russia, has become our principal foreign source. In 1914, Brazil furnished two-fifths of the imports and more than India or Russia, and in 1915 over nine-tenths. The manganese ores of Brazil are consequently of more than usual interest to us at this time. The Manganese-Ore Distbicts of Brazil , The manganese mining industry of Brazil dates from the year 1894, and since that year the total production has been over 3,000,000 tons. With the exception of a small quantity produced in the State of Bahia from deposits west of the City of Bahia, this output has come from the

Jan 1, 1917

A. L. FIELD, Cleveland, Ohio (written discussion*.-)In equation 2, B is used to denote the ratio of bases (lime plus magnesia ) to acids (alumina plus silica) it being stated that this ratio gives more concordant results than the ratio either of bases to silica, or of lime to silica. I find, however, that, by making use of the ratio between true CaO (CaO by analysis minus sulfur calculated to CaS) to acids, equation 2 may he modified to agree with the observed facts within an average error of 0.0041 per cent. sulfur instead of 0.0045 per cent. The revised equation is s = 0.072 + 0.01835 - 0.00061 (t - 1400) - 0.0403B where B is the ratio of true CaO to Al2O3 Plus Si02, and the other symbols have the same significance as before. I would not appear to stress too much this slightly better agreement of the revised equation with the observed sulfur. Yet I believe that the use of this ratio of true CaO to acids is to be preferred, if for no other reason than the fact that it is more in harmony with the previous findings of the Bureau of Mines with regard to the effect of magnesia and other impurities on slag viscosity.1

Jan 12, 1919

[Appalachian Petroleum Section: James M. Kimbrough, chairman; John H. Adair, first vice-chairman; Walter L. Diddle, second vice-chairman; Byron M. Finch, secretary-treasurer. Balcones Section: Robert V. West, chairman; Frank W. Jessen, first vice-chairman; Zane V. Morgan, second vice-chairman; Brantley G. Myers, secretary-treasurer. Meets 4th Wed., Petroleum Club (San Antonio) or Terrace Dining Room (Austin) Billings Petroleum Section: Charles B. Evans, chairman; L. M. Lyon, first vice-chairman; Gordon E. Lansford, second vice-chairman; Fred. A. Haddenhorst, secretary-treasurer. Meets 3rd Tues., Billings Petroleum Club. California Coastal Section: Robert L. Richardson, chairman; David K. Hayward, first vice-chairman; W. Jackson Bransford, secretary; William N. Harwell, treasurer. Meets 3rd Thurs., Pierport Inn, Ventura. Caracas Petroleum Section: Edwin A. Moo re, chairman;. Jose Cirigliano, vice-chairman; Frank B. Davis, secretary; Robert N. Powers, treasurer. Meets 2nd Tues., Caracas Theater Club.]

Jan 1, 1961

By H. Reiss

The zone-melting process in which redistribution of solute in a solid bar is effected by the passage of a molten zone is considered mathematically. Simple approximate techniques are developed for computing the manner in which redistribution occurs as molten zones continue to pass. The introduction of the "zone-melting flux" provides valuable insight into the nature of the phenomenon, as well as a central mathematical theme in terms of which the process can be discussed. Curves are presented for typical and general cases. AN extremely efficient purification technique was described recently by Pfann.1 Known as zone refining, it consisted of the slow passage of a series of molten zones, produced by external heaters, through a long solid ingot of the substance to be purified. Equations were given for the maximum separation attainable, and a numerical computation method was described for obtaining the concentration in the ingot after a given number of zone passes. The present paper offers an interpretation of the physical nature of zone melting in terms of a transport process in which formal diffusive and convec-tive flows occur. This interpretation, in addition to providing insight into the physical nature of the process, also provides a basis for its mathematical treatment. From this basis, useful equations are developed for the solute concentration in the ingot as a function of the number of zone passes. These are applicable particularly where large numbers of passes are necessary and where numerical computational methods might be laborious. All considerations will deal with the limiting form of the process, described by Pfann,' in which: 1—Diffusion in the solid is considered negligible. 2—Mixing in the liquid is supposed, complete. 3—The distribution coefficient, k, measuring the ratio (at equilibrium) of solute concentration in the solid to that in the liquid is supposed, independent of composition. In general, this paper will be concerned with a bar of uniform (actually unit) cross section, ex- tending in the x-direction from x = 0 to x = L, where L may be infinite. Zone melting is a transport process, and as such must admit of interpretation in terms of a flux of transported quantity. However, the progress of operation is not measured in units of time but in terms of the number of completed zone passes. Consequently, the zone-melting flux will be a rate, per unit pass, rather than per unit time. The description of this flux and the examination of its properties is given below. Zone-Melting Flux Fig. 1 represents an intermediate stage of the nth pass. Concentration C is plotted against distance, x. The molten zone with its back at the point x and front at x + I moves in the direction indicated by the arrow. As it moves, it leaves behind the nth

Jan 1, 1955

By H. P. Allen

The Achotla mill of the Cia. Minera de Penoles is of special interest in that it is one of the few still employing the chloridization process. The ore milled comes from the Suriana mine and consists of both oxide and sulphide, from neither of which can the gold or silver be extracted by a simple cyanide treatment. The mill was built in 1927, after several years of study in the development of the process1, and operated from December of that year until October 1930. Operations were resumed in February 1935, and have been continuous since, treating approximately 5000 tons per month, while the mine is getting up to full tonnage of 6500 tons per month. Transportation of Supplies The property is in Guerrero, Mexico, some 175 km. by airline southwest of Mexico City. The mill site is at Temisco (Fig. I), 8 km. from the Balsas River and 4.8 km. by aerial tram from the mine. At Temisco are the general office, shops, warehouses and laboratory serving both mine and mill. Here sufficient water is available either from a small creek running through the property or from the Balsas River, also convenient and ample storage space for the tailings. With the exception of the mine timber and a few native building materials which may be obtained locally, all supplies for both mine and mill are brought down the Balsas River from Balsas, the terminal of a Mexican National Railway line out of Mexico City. The method of transportation on the river is interesting. Rapids with very fast water are numerous, prohibiting the use of mechanically propelled boats. Those used resemble large, flat-bottomed rowboats some 22 ft. long by 8 ft. wide (Fig. 2). They are manned by two oarsmen sitting in the bow, each handling one large oar, and usually another man in the stern who, with his oar, assists in steering the boat, especially when the rapids are run. From 3 to 7 tons of freight are carried on each boat, depending upon the stage of the water, which governs also the time

Jan 1, 1939

By H. P. Allen

The Achotla mill of the Cia. Minera de Penoles is of special interest in that it is one of the few still employing the chloridization process. The ore milled comes from the Suriana mine and consists of both oxide and sulphide, from neither of which can the gold or silver be extracted by a simple cyanide treatment. The mill was built in 1927, after several years of study in the development of the process1, and operated from December of that year until October 1930. Operations were resumed in February 1935, and have been continuous since, treating approximately 5000 tons per month, while the mine is getting up to full tonnage of 6500 tons per month. Transportation of Supplies The property is in Guerrero, Mexico, some 175 km. by airline southwest of Mexico City. The mill site is at Temisco (Fig. I), 8 km. from the Balsas River and 4.8 km. by aerial tram from the mine. At Temisco are the general office, shops, warehouses and laboratory serving both mine and mill. Here sufficient water is available either from a small creek running through the property or from the Balsas River, also convenient and ample storage space for the tailings. With the exception of the mine timber and a few native building materials which may be obtained locally, all supplies for both mine and mill are brought down the Balsas River from Balsas, the terminal of a Mexican National Railway line out of Mexico City. The method of transportation on the river is interesting. Rapids with very fast water are numerous, prohibiting the use of mechanically propelled boats. Those used resemble large, flat-bottomed rowboats some 22 ft. long by 8 ft. wide (Fig. 2). They are manned by two oarsmen sitting in the bow, each handling one large oar, and usually another man in the stern who, with his oar, assists in steering the boat, especially when the rapids are run. From 3 to 7 tons of freight are carried on each boat, depending upon the stage of the water, which governs also the time

Jan 1, 1939

By R. Simon

Maurer's review1 summarizes quite thoroughly the various theoretical developments and experimental findings that contribute to the knowledge of rock mechanics in drilling. This discussion supplements Maurer's paper by including generalities that either have been incompletely treated in his review, or formulated from new considerations by the discussant and from hitherto unpublished experimental data. These generalities are associated with the observation that all mechanical methods of rock drilling essentially comprise a sequence of indexed, concentrated loadings on the bottom of the hole being drilled. A concentrated loading is defined as a mechanical loading by a tool that contacts at any one time only a small fraction of the surface of the body being worked and penetrates only a small fraction of the body thickness. The surface in this application is the bottom of the hole. The deformation of ductile material by concentrated loading has received considerable attention because of the importance of indentation tests for measuring hardness of metals. Some of the generalities involved in concentrated loadings on ductile materials apply also to concentrated loadings on more brittle materials such as rock. Among these generalities is the applicability of similitude considerations to scaling laws and to force-displacement and energy-volume characteristics for tools of various shapes, as summarized by Tabor.2 Vork on the flow and fracture behaviors of glass loaded with pyramidal and spherical indenters is reviewed by Douglas.3 Maurer refers to the work of Evans and Murrell on wedge loadings on coal and on mixtures of pitch and sand. It is evident from Maurer's paper that a considerable portion of the studies of rock mechanics in drilling, such as crater formation experiments, has involved concentrated loadings on rock.

Jan 1, 1967

By G. M. Ritcey, A. W. Ashbrook, E. G. Joe

Solvent-in-pulp extraction of uranium from acid leach slurries of flotation concentrates is described. Solvent extraction was carried out in a 10 in. diam pulsed-plate column. The extractant was a 0.1 M tertiary amine (General Mills Alamine 336) in a kerosene diluent containing isodecanol as an emulsion suppressor and third phase inhibitor. In pilot plant operation at the Beaverlodge mill of Eldorado Mining and Refining, Ltd., slurry feeds containing up to 35% solids were processed at a throughput rate of 13.5 tpd of dry feed. Uranium extraction obtained was as high as 98%. Amine losses to the raffinate were approximately 0.07 lb amine per ton of original dry feed. The Beaverlodge Mill of Eldorado Mining and Refining employs a carbonate leach for the extraction of uranium from the ore.',' Prior to carbonate leaching sulphides are removed by flotation to reduce reagent consumption. This flotation concentrate containing some uranium is then leached with sulfuric acid, filtered, and the uranium precipitated with magnesium hydroxide. The precipitate is then returned to the carbonate leach circuit. Although the tonnage of flotation concentrate treated by acid leaching is low, reagent costs are relatively high. Also maintenance costs on the acid leach filtration circuit were beginning to show significant increases. The choice of solvent extraction-in-pulp as an alternative to the present circuit was influenced by the following factors: 1) elimination of acid pulp filtration, 2) increase in overall recovery because of soluble loss elimination, 3) reduction of reagent cost (mill sodium carbonate solutions are used as strip), and 4) production of higher purity precipitate at the mill site. Because a solvent-in-pulp process would be difficult to test on a bench-scale, the Eldorado experience in the use of pulse columns,3-5 mixer-settlers,6 and solvent-in-pulp operations in particular7 was applied directly to the initial pilot plant carried out during the summer of 1965 and followed by the second pilot plant in 1966. The latter work, which is presented in this paper, determined the following factors for anticipated scale-up to plant production: 1) overall uranium recovery by this process, 2) tonnage throughput expected, and 3) solvent losses anticipated for the full-scale plant. The equipment used consisted of a 10 in. diam extraction pulse column and three mixer-settlers for scrubbing, stripping, and acid equilibration. All equipment was designed by Eldorado and much of the pilot plant was built at the mill site. PROCESS CHEMISTRY Feed to the acid leach plant consists of a sulphide flotation1,2 concentrate containing mainly pyrite, uranium and graphite, with a particle size of 90%

Jan 1, 1968

By H. Ries

Introduction BODIES of pyritic ore in schistose rocks have long been known in different parts of the world. The several occurrences resemble each other in being usually of more or less lenticular shape, inclosed in walls of schist or gneiss, and carrying pyrite, chalcopyrite, and pyrrhotite in varying proportions, as the chief ore minerals. The deposits differ, however, in that sometimes one, sometimes another, of the three sulphides mentioned may predominate. The orebodies in general are more or less closely conformable with the schistosity of the wall rocks, although in cases this may not hold true. The boundaries may or may not be sharp, but in general are fairly distinct. The lenses vary in size, may occur singly, or in groups; in the latter case they may be in line, overlapping, or occasionally parallel. Pinching and swelling of individual lenses is not uncommon. The genesis of these pyritic bodies has provoked much discussion, and they have been variously classed as sedimentary, igneous intrusions, hydrothermal replacements or impregnations, etc. Indeed, in some cases different geologists have most positively assigned widely diverse origins to the same deposit, as in the case of Rio Tinto 1 and Rammelsberg.2 It is true that in the light of modern criteria, we have been able in a few cases to decipher clearly the process of origin and so there seems no doubt now regarding the genesis of the deposits at Rio Tinto, Spain;3 Rammelsberg, Germany ;4 Ducktown, Tenn.;5 southwestern Virginia;6

Jan 8, 1917

By Robert E. Green, Edmund G. Henneke

The velocities of the three possible modes of elastic wave propagation have been calculated for single-crystal copper at 1-deg intervals throughout the standard stereographic triangle. The results are presented as isospeed contours. The deviations between the direction of maximum energy flux and specimen axes are presented as arrows drawn on stereographic projections. The orientation factors associated with dislocation motion on the primary slip system only and also on all twelve slip systems have been calculated for all three elastic waves. The orientation factors associated with a longitudinal standing wave resonant specimen have also been calculated. The results of these calculatians are compared with earlier calculations on aluminum crystals and the relevance of these results to ultrasonic attenuation measurements is discussed. GREEN and inton' have recently outlined a procedure for determining the orientation factors for dislocation damping in single crystals. The stress field associated with a given mode of elastic wave propagation is resolved on the slip systems in which it is assumed the dislocations are vibrating due to this stress field. Since, at room temperature, slip has been observed on a limited number of systems in fcc and hcp metal crystals, it is assumed that these systems are the appropriate ones to use for the resolving process. This procedure has been applied to aluminum crystals by Hinton and hreen' and to zinc crystals by Henneke and Green.3 In the present paper the modes of elastic wave propagation have been calculated for copper crystals possessing orientations at 1-deg intervals throughout the standard stereographic triangle. The deviation of the energy flux vectors from the specimen axis have been calculated for each mode. The stress field associated with each mode has been resolved on the (111)(110) slip systems in order to determine the orientation factors for dislocation damping. MODES OF ELASTIC WAVE PROPAGATION For any direction in a linear homogeneous aniso-tropic material three elastic modes can be propagated. In general, each of these modes has both longitudinal and transverse particle displacements associated with it and thus are not "pure" modes. However, there is always one component which is greatest in either a longitudinal or transverse direction and hence the mode may be referred to as quasi-longitudinal or quasi-transverse, accordingly. For certain axes of high symmetry pure modes may be propagated. The phenomenon of refraction of the energy flux vector is also exhibited by elastic wave propagation in ani-

Jan 1, 1969

By R. E. H. Pomeroy

The statistical data given in this paper are taken from the actual performance of the No. 2 reverberatory furnace of the Nevada Consolidated Copper Co., Mc Gill, Nev., for a period of four months, from April to July, 1914. The principal furnace dimensions are: Length, 132 ft. % in. at the skim-line level; average width, 18 ft., 10 in. at the skim-line level. The roof is 7 ft. above the skim-line level at the firing end. The average area in the uptake flue is 36 sq. ft., and the area above the skim line under the vulcatory arch is 24.3 sq. ft. Fig. 1 shows a diagrammatic plan and sectional elevation of this furnace. The furnace is equipped with two 400-b.h.p. Stirling waste-heat boilers set in parallel. California crude petroleum is used for fuel and has a specific gravity of 16.5" B., flashing in open test at 199" F. The oil is heated to about 200" F. in a steam heater before going to the burners, requiring about 9 b.h.p. per furnace day to heat it from the line temperature to the burning temperature. Measurement is made by oil meter, checked daily by reservoir readings. The oil is fed by gravity (34 Ib. static pressure) to seven low-pressure blast burners of the Steptoe type (Fig. 2). The blast for these burners is supplied at about 40-08. pressure by a motor-driven Connersville blower of 42 cu. ft. capacity per revolution. The air supplied by the blower through the burners for atomizing the oil amounts to about 10 per cent. of the theoretical air necessary for complete combustion of the oil. The remainder of the necessary air enters the furnace through the burner openings in the firing wall. No checker holes are provided, and the charge holes in the roof are equipped with slide gates and " dog houses" to minimize air leakages. The combustion of the fuel, as shown by gas analysis at the front of the furnace, is practically complete, the carbon monoxide in the gases being less than 0.5 per cent. The draft at the firing end of the furnace is 0.18 in. of water. At the throat or in the uptake above the "verb" the draft is about 1 in.

Jan 1, 1915

By J. T. McGrath, G. B. Craig

Tensile tests on aluminum single crystals, grown from the melt, revealed that the yield stress of these crystals was raised as the amount of striation substructure was increased. The number of striation-type boundaries was varied by changing the rate of growth of the crystals. The results are interpreted in terns of a blockage of dislocation movement by striation boundaries. IMPERFECTIONS are known to have a large effect on the mechanical properties of metals. Teghtsoonian and Chalmers 1 studied a block or striation-type substructure in tin and found that the formation of the striations depended upon rate of crystal growth and orientation. They found that the angular misorienta-tion between neighboring striations varied between 15 min and 5 deg. Atwater and chalmers2 investigated striations in tin and lead, and found that purity was another factor which influenced their formation. There was a limiting purity above which the striations were straight and regular. There is little information in the literature concerning how substructures in melt-grown crystals affect strength. washburn3 introduced small-angle boundaries into zinc by a shear process and found a subsequent increase in yield stress which he attributed to the presence of these boundaries. Lauriente and pond4 found the critical resolved shear stress in aluminium crystals grown from the melt, to increase with etch pit density. The etch pits were used as an indication of growth imperfections. In Hibbard's5 work on lineage structure in copper crystals grown from the melt, he tested one crystal containing interlineage boundaries, and found a 50 pct increase in the critical resolved shear stress. Fleischer and Chalmers6, in a recent paper, mention briefly that, in the production of aluminium crystals from the melt when the growth rate was raised from 1.2 to 18 cm per hr, the yield stress increased 50 pct. They felt that the increase in yield stress with rate of growth might be attributed to a higher concentration of impurities in the gage length and also might be affected by the increased number of small-angle boundaries (striations). The present investigation was undertaken to determine the effect of the striation substructure on the strength of aluminum single crystals. EXPERIMENTAL PROCEDURE Aluminum single crystals of 99.9992 purity were grown from the melt using a technique described by Chalmers.7 The growth rate was varied from 0.6 to 20 mm per min, thus producing crystals with varying numbers of striation boundaries. To avoid prior deformation, the crystals were chemically cut to the desired length with a saturated solution of cupric chloride. They were then annealed at 580°C for 24 hr and allowed to cool slowly in the furnace. Finally, the specimens were electropolished using an electrolyte of 5 parts ethyl alcohol and 1 part perchloric acid. Each specimen was approximately 5 in. in length and 0.125 sq in. in cross section. The average orientations of the 5 series of crystals studied are shown on the stereographic projection of Fig. 1. The crystals were strained approximately 2 pct in a soft tensile machine. The strain sensitivity was 3.6 x 10-4 cm over a 5-cm gage length.

Jan 1, 1960

LIST OF THE MEETINGS OF THE INSTITUTE AND THEIR LOCALITIES FROM ITS ORGANIZATION Transactions Number Place Date Vol Page 1 Wilkes-Barre, Pa May, ?71 1 3 2 Bethlehem, Pa August, ?71 1 10 3 Troy, N Y November, ?71 1 13 4 Philadelphia, Pa February, ?72 1 17 5 New York, N Y May, ?72 1 20 6 Pittsburg, Pa October, ?72 1 25 7 Boston, Mass February, ?73 1 28 8 Philadelphia, Pa May, ?73 2 3 9 Easton, Pa October, ?73 2 7 10 New York, N Y February, ?74 211 11 St Louis, Mo May, ?74 3 3 12 Hazleton, Pa October, ?74 3 8 13 New Haven, Conn February, ?75 3 15 14 Dover, N J May, ?75 4 3 15 Cleveland, 0 October, ?75 4 9 16 Washington, D C February, ?76 4 18 17 Philadelphia, Pat June, ?76, 5 3 18 Philadelphia, Pa October, ?76 5 9 19 New York, N Y February, ?77 5 27 20 Wilkes-Barre, Pa May, ?77 6 3 21 Amenia, N Y October, ?77 6 10 22 Philadelphia, Pa. May, ?78 6 18 23 Chattanooga, Tenn May ?78 7 3 24 Lake George, N.Y. Oct. ?78 7 25 Baltimore, Md. Feb. ?79 7

Jan 1, 1923