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By P. R. Hines

Diphenyl guanidine is used as an accelerator in vulcanizing rubber. Other rubber accelerators are also flotation collectors, e.g., dithiocarbamate, thiazole, and the xanthates. Urea and its derivatives are good flotation collectors,' and guanidine is the nitrogen analog of urea. These two characteristics suggested testing diphenyl guanidine as a flotation collector. Diphenyl guanidine has been tried in flotation work previously, but references give no details.'.' Bunker Hill Co. of Kellogg, Idaho, supplied the ore sample used in these tests. The sample was typical of the ore milled in 1949 and contained 4.52 pct Pb and 1.08 pct Zn. A flotation test made by the author with potassium ethyl xanthate as a collector, the same flotation reagents, and the same grind employed in regular Bunker Hill mill practice was the standard for comparing results with the compounds and derivatives of guanidine. The compounds and derivatives of guanidine are only fair collectors of galena. However, if potassium ethyl xanthate were used in the galena float, its presence later in the sphalerite float would make it impossible to introduce another type of collector and be certain which collector contributed the result, so the compound or derivative of guanidine under test was used in both the galena and sphalerite floats in Table I. On both the galena and the sphalerite in the Bunker Hill ore, the depressing action of sodium cyanide and zinc sulfate is much greater with collectors of the guanidine type than with the xanthates. If the depressing agents are left out when a guanidine type of collector is used, but a pH of 8.6 is maintained with either lime or sodium carbonate, the depressing action is approximately the same as in the standard xanthate test. Consequently the depressing agents, sodium cyanide and zinc, were not used in the galena floats of tests 1226, 1228, and 1286, Table I, so the galena floats would be on a comparable basis with the standard xanthate test. As shown by the tests in Table I, the loss in the tail when diphenyl guanidine is used is 8.3 to 9.3 pct, as compared to 17.6 pct with potassium ethyl xanthate, or a recovery of 2.3 to 2.0 lb more zinc. Concentrate grade is 13.7 pct with potassium ethyl xanthate and increases to 22.9 and 31.8 pct with diphenyl guanidine, showing a much higher selectivity. Unless the Barrett's No. 4 coal tar creosote is considered to be a collector, no collector other than diphenyl guanidine is present in the tests, Nos. 1226, 1228, and 1286. Tests 1218, 1253, and 1263 in Table I1 were run to check the effect of the depressing agents used in flotation of the galena (sodium cyanide and zinc sulfate) upon subsequent flotation of sphalerite by diphenyl guanidine. Urea was substituted for potassium ethyl xanthate in tests 1253 and 1263 because it is a good collector for galena and a poor one for sphalerite, and the effect of the depressing agents is more marked with urea than with potassium ethyl xanthate. Test 1205 checks the effect of urea alone without any depressing agents when used for the flotation of galena. Table I1 shows that diphenyl guanidine as a collector for sphalerite is compatible with other flotation reagents used in floating galena. When the depressing agents sodium cyanide and zinc sulfate are used in the galena flotation step, the amount of lime in the sphalerite flotation must be increased (compare 1253 and 1263). Table I11 gives some of the compounds of guanidine which were tested and compared. Some of the compounds and derivatives of guanidine* are solu- * Diphenyl guanidine is in commercial production and is available on the market. ble in water and others are not. Those tested were fed dry. When an organic collector is promising, its compounds and derivatives should be tested and their comparative collecting power recorded. Eventually, in this way, it may be possible to discover some of the essential chemical characteristics of a good collector. For example, it is interesting to compare the diphenyl and dibutyl derivatives in Table IV with those of urea in Table 111, Example 3, U. S. Patent 2,664,198. All are good collectors. Table I shows some of the effects of lime on recovery of sphalerite by diphenyl guanidine. Table V

Jan 1, 1960

By Y. C. Liu, W. R. Hibbard

Based on pole figure data and microstructural observations, the re-crystallization orientation found in a copper strip previously cold-rolled 99.5 pct from a single crystal with an initial (110) [112] orientation may be described as: a 30° rotation, clockwise and counterclockwise, about octahedral poles of the cold-rolled texture such that all four poles function at a low annealing temperature (400°C) and only three of them function at high temperatures (500' to 1050°C). The relative intensity of deformation stresses on various slip planes can be correlated with the choice of poles affecting the rotations found in the recrystallized orientations. ROTATIONAL reorientation about the poles of densely packed crystallographic planes is an important characteristic of secondary recrystallization" in face-centered cubic metals. Some evidence has also been interpreted to show that such a rotational relationship also exists in the primary recrys-tallization process." In an investigation of the crystallographic relationships in the recrystallization process, several aspects of the experimental procedure should be considered. With cold-rolled poly-crystalline metal as the initial material, pole figure analysis leaves many ambiguities as to whether or not simple indices can be assigned to adequately represent its orientation. The rolling texture does not consist of a single texture. Minor deformation textures are usually present. In all cases, twin textures are present. The presence of both minor and twin deformation textures influences the recrystallization texture. In order to study the recrystallization mechanism, a more precise knowledge of the history of the material before recrystallization is necessary. Tensile deformation of a single crystal usually produces residual stresses which are concentrated at slip bands. Since the exact nature and orientation of these potential nuclei sites cannot be experimentally evaluated, the tensile deformation of a single crystal was not considered adequate. From the work done by Barrett and Steadman6 n copper, it appeared that cold rolling of a single-crystal specimen with an initial (110) [112] orientation would yield a specimen which fulfills the conditions required for the present investigation, namely a highly preferred single rolling texture with relatively homogeneous stress distribution. By investigating the recrystallization texture of this material, additional details of reorientation during recrystallization might be obtained. The purpose of this paper is to describe such an investigation. Experimental Procedure Copper used in the present investigation was cathode sheet with a purity of 99.94 to 99.97 pct. A copper single-crystal specimen, 1.15x0.8720x0.5322 in. (thickness) was cut from a cylindrical crystal which was grown in a Bridgman furnace.' The cutting was accomplished on a horizontal milling machine operated at a very slow speed to obtain a (110) plane in the rolling plane and a [112] direction in the rolling direction. The disturbed surface was removed by electrolytic polishing in dilute orthophos-phoric acid (H3PO4) with a specific gravity of 1.14, and a current density of about 1 to 2 amp per sq cm. A 3-min polishing was needed to eliminate the disturbed surface causing Debye-Scherrer rings and a 45-min polish produced a surface which yielded sharp rounded Laue spots. It was estimated that about 1/16 in. of metal was removed from the original cut surface. The final orientation of this specimen before rolling was about 2" from the (110) plane in the rolling plane with a [112] direction aligned in the rolling direction. It is possible that a small amount of strain was still present in the cube despite the fact that the Laue spots seemed very sharp. It is doubtful, however, that this would be an important factor after a subsequent rolling reduction of 99.5 pct. This copper specimen was rolled on a laboratory rolling mill with two highly polished rolls 37/8 in. in diam in the following manner: 0.010 in. per pass to 50 pct reduction, 0.005 in. to 70 pct, 0.002 in. to 85 pct and, finally, 0.001 in. per pass to a 99.5 pct reduction. Specimens, about 1 in. sq with the rolling edges intact, were cut from the rolled strip by the electrolytic method previously described, after the strip

Jan 1, 1954

By Hsun Hu, R. S. Cline

The effect of rate of deformation on texture formatiotz has been studied with cube-oriented single crystals of 3 pct Si-Fe, compressed 80 pct at two widely different rates. Compression at a low rate (crosshead speed, 0.1 ipm) produces large orienta-tional changes and coarse deformation bands, whereas little change in orientation is observed, particularly at the surface of the crystal, when compressed at a high rate (impact velocity, -500 ips). Thin plates of mechanical twins are produced at the high rate, but their contribution to the deformation texture is practically negligible. The crystal deformed at a high rate has a lower hardness and a more uniform dislocation structure and is morc resistant to recrystallization. There is also a dilference in the recrystallization texture between the slowly and rapidly compressed crystals. These results indicate that plastic flow is less "turbulent" at high strain rates than at low strain ratcs. THE behavior of metals under high-speed deformation has been the subject of wide study in the past decade. Most of these works were concerned with mechanical properties during testing at high rates of loading, the resulting structures, and the theoretical aspects of the observed phenomena. The present status of knowledge and research activities in this field has been documented in recent publications.17' There is, however, very little information available in the literature concerning the effect of high-speed deformation on the texture formation in metals. The possibility of an effect of rate of deformation on texture is suggested by the microstructural differences observed between normally and very rapidly deformed samples. Aside from the strong tendency for twinning (and phase transformation in certain metals) in high-speed deformation, the slip-line patterns and the dislocation configurations show marked differences as a result of large differences in the strain rates. For instance, in mild steel in static and dynamic compression, Campbell and co-worker3,4 observed that coarse slip occurred on several systems in specimens deformed at normal strain rates, whereas specimens deformed rapidly showed only fine slip on fewer systems, the slip-line pattern being very similar to that in specimens deformed at low temperatures and at normal strain rates. The absence of multiple slip was also noticed by Dieter5 in shock-loaded nickel. The dislocation configuration in an iron shock loaded at 70 kbar pressure was shown by Leslie et al.6 to be similar to that produced by light rolling at low temperature—the dislocation lines were mostly straight and uniformly distributed. These observations suggest that the mode of deformation at high strain rates is notably different from that at low strain rates. Consequently, the nature and the extent of lattice reori-entation during deformation should also be different. This investigation was undertaken to test this idea. MATERIAL AND METHOD The remaining part of a single-crystal ingot of a 3 pct Si-Fe alloy,* prepared for an earlier investi- gation,1 was used in the present studies. Specimens 9/16 in. square and 0.365 in. thick were cut from this ingot with (001) planes parallel to the square surfaces, and the [loo] and 1010] directions parallel to the edges of the specimen, within ±1 deg. These machined crystals were etched to remove distorted metal, and annealed at 1300°C for 24 hr in a purified helium atmosphere. The annealed crystals had a hardness of 161 Dph. The crystal orientation was rechecked with X-ray back-reflection techniques. In high-strain rate experiments the crystal was compressed in a Dynapak machine to approximately 80 pct reduction in thickness (from 0.360 to 0.070 in. with an impact velocity of -500 ips). Thus, the duration of loading was approximately 6x 10-4 sec, and the average strain rate was about 103 sec-1. As a result of this rapid, severe deformation, the specimen became a roughly circular thin disc. The original shape of the crystal, however, could still be recognized by a diffusely outlined square area in the center of the disc. For compression at a low speed, the crystal was deformed in a universal testing machine at a rate of crosshead movement of 0.1 ipm. Compared with

Jan 1, 1965

By Jr. Ensign C. O., J. W. Trammell

The two principal methods of room-and-pillar mining practiced at White Pine make it important to predict variations in the thickness and rock types of a stratum called the upper sandstone. In full column mining, the nearly barren upper sandstone occurs between two ore horizons (called upper shale and parting shale, respectively), and is mined as a part of the ore column. In general, full column mining is not practiced where excessive upper sandstone thickness causes full column ore grade to be substantially lower than parting shale ore. In parting shale mining, the upper sandstone forms the mine roof. Difficulty in predicting upper sandstone character arose because the extent and directions of its variations were not apparent in the property's diamond drilling, in which holes are spaced on 1000-ft centers. A study of sedimentary features, undertaken to improve the predictability of trends in the upper sandstone, led to an interpretation of the sedimentary environment in which the sandstone was deposited. Ripple marks, mud cracks, and cross-bedding, as well as other, less well known features, such as channel casts, flute casts, and current crescent casts, were mapped or recorded. These data, coupled with the knowledge of regional facies changes gained from studying drill core, show that the upper sandstone was deposited by a series of streams flowing northeastward over the underlying parting shale. Deposition of the lowermost bed of the upper sandstone, near the ancient shoreline, was locally preceded by erosion of the parting shale, and the greatest thicknesses of sandstone are found in channels scoured out of the parting shale. Awareness of the rather strong linear trends in the upper sandstone makes it possible to project continuous areas of thick upper sandstone through apparently isolated "highs" in the upper sandstone thickness contour map, which is based on drill-hole information. Since local exceptions to general stratigraphic trends exist, a method was also needed for esti- mating, in detail, the thickness and degree of shali-ness of upper sandstone forming the roof in active parting shale mining areas. Because locally the parting shale was thinned by erosion during the deposition of upper sandstone, parting shale thickness is inversely proportional to upper sandstone thickness. Utilizing information taken from short drill holes into the roof, curves were constructed for the correlation of parting shale thickness (measurable in the mine) with the upper sandstone total thickness, as well as the thickness of its basal member. The White Pine mine is situated in Ontonagon County, approximately 6 miles south of Lake Superior, in the Upper Peninsula of Michigan (Fig. 1). The orebody mined by White Pine Copper Co. occurs in the lowermost 20 to 25 ft of the Nonesuch formation, a series of middle- to upper-Keweenawan shales, siltstones, and sandstones. This paper discusses the solution of problems in ore grade control and mine planning arising from the presence of an essentially barren sandstone stratum within the orebody. THE OREBODY AND MINING TYPES A brief description of the orebody and mining methods is necessary to show how the upper sandstone affects mine planning and grade control. In the mine, the cupriferous zone of the Nonesuch shale is divided into three major portions, as shown in Fig. 2. Lying conformably on top of the Copper Harbor formation (or "lower sandstone") is the lower part of the orebody, the parting shale. The parting shale is overlain by the upper sandstone. On top of the upper sandstone, the basal portion (8 ft) of the upper shale is quite similar to the parting shale. Both the upper shale and parting shale are divided into a number of smaller stratigraphic units, each having a characteristic copper content, as shown by the histogram of Fig. 2. Other authors1'3 have dealt in detail with the stratigraphy and mineralogy of the orebody. However, in this article, we are concerned with the major subdivisions only, and it will suffice to note the low copper content of the upper sandstone, shown in Fig. 2. At present, two types of room-and-pillar openings are created in the orebody. In one type, called parting

Jan 1, 1964

By N. G. W. Cook

The conventional cyclic system of deep-level mining by drilling and blasting gives rise to an inadequate degree of stope sorting when mining thin reefs. This results in poor utilization of the capital facilities of a mine in the form of shafts, haulages, airways, and associated equipment. Continuous and controlled removal of the thin gold-bearing portion of the reefs would permit better stope sorting and hence greater utilization of capital facilities. Results of experiments to develop hard-rock cutting machines for mining are reported and the benefits which could be derived from their use are discussed. Mining, from exploration through refining, is essentially a process of sorting in which the payable mineral, or metal, is progressively separated from the other constituents of the earth's crust with which it was originally associated. This takes place more distinctly at each step of the operation. What is it that determines the degree to which sorting should be carried out at each of the several steps comprising a whole mining operation? The formal answer is that degree of sorting at each step which results in the lowest overall cost for the complete separation. In practice, individual steps are chosen from those available in current technology, each of which effects a degree of sorting such that the quantity of material which must be sorted in the succeeding step is economically acceptable. It follows that any new technological development has repercussions throughout the whole mining operation and, more important, that the solution to excessive costs in any one step of the operation may lie not in improving the costly operation itself so much as in increasing the degree of sorting preceeding that operation. This concept, particularly in relation to deep-level mining of thin, tabular gold-bearing reefs in South Africa, is discussed here, and the most recent results achieved in the development of hard-rock cutting machines for stoping more selectively than is possible with explosives are presented. Deep-Level Mining Deep-level mining involves operations which are either not encountered, or are of only trivial importance, in near-surface mining. Near-surface, the major operations are those of rock breaking, transport, and milling. In deep-level mining, hoisting, environmental control, and strata control assume major importance. Some idea of the relative magnitude of these operations may be gained by comparing the separate amounts of energy which are required, or which must be controlled, to effect the various operations when, say, mining a tabular deposit 40 in. thick at 8000 ft below surface, Table 1. It is true that the costs of handling a given quantity of energy are not the same for each operation. Nevertheless, Table 1 does emphasize the fact that the operations of hoisting, strata control, and environmental control are of unique and major significance in deep-level mining. In particular, hoisting and environmental control place a heavy load on the reticulation system of the mine—the shafts, haulages, and airways. Typically, a new, deep gold mine with an annual revenue of about $35 million requires a total capital expenditure of about $140 million of which some $100 million is invested in developing and equipping this reticulation system. The ratio between annual turnover and capital invested of about one-quarter is exceptionally low, and it typifies the poor utilization of capital by the current technology of mining hard rock at depth. The average thickness of the reefs in the new South African goldfields varies from 10 to 30 in.,l and even in the thicker reefs the gold is often confined within a small fraction of the nominal thickness. Nevertheless, it is universal practice to mine these reefs at a stope width of about 40 in. or more, so that the quantity of rock broken in the stopes and hoisted out of the mine is between two and ten times the quantity of rock actually carrying a significant amount of gold. The reason for the adoption of such excessive stope widths is to be found in the method of rock breaking by drilling and blasting. The only free surface to which a blast hole can break is the stope face. It follows2 that each hole cannot have a burden in excess of the height of the free face if it is to break satisfactorily. To

Jan 1, 1971

By R. E. Zinkham

An investigation has been conducted to compare short transverse and longitudinal fatigue and fracture properties in 4.25-in.-thick, high strength aluminum alloy plates. One plate was produced using standard rolling techniques while the other was pre.forged before rolling. Little difference was shown in fatigue strength of longitudinal specimens taken from mid-thickness of the plate. Howeuer, in the short transverse orientation fatigue strengths at 107 cycles were about 25 and 50 pct less, respectively, for the preforged and standard rolled plate. Differences in fatigue strengths were attributed to grain size and shape as well US orientation of constituents. Fatigue crack propagation rates and fracture toughness were compared at three different stress intensity (K) levels, using a constant compliance, double cantilever, wedge-shaped specimen. In a given plate, comparable fatigue crack Propagation rates were observed in the longitudinal (i9W) and short transverse (TW) orientations. Somezuhat gveater rates were observed in the short transzerse (TR) orientation. The preforged plute gave a lower rate for all three directions. Considerable secondary cracking developed, at times, over portions of the fatigue crack in both plates, particularly at the lower stress intensity levels in the short transverse specimens. Micro structure revealed constituent stringers as possible causes of the crack branching. Fracture toughness was considerably less in both plates in the short transuerse orientation. It is concluded that preforging not only improved directional tensile properties but also the fatigue and fracture properties in general. On occasion, aluminum plates have been milled away for hinges or bolted connections and stressed through the thickness or short transverse direction. Little or no information is available concerning fatigue characteristics or fracture toughness in this loading orientation in aluminum plate, or of the effect of fabrication on these properties. It was the intent of this project to examine, develop, and apply a unique specimen that has been advocated by others to study the fatigue characteristics and fracture toughness of two differently fabricated high strength aluminum plates. Linear elastic fracture mechanics criteria may be applied to the specimen so that the fatigue crack propagation rate and fracture toughness data may be of use for design or inspection applications. Fatigue characteristics are generally measured in the longtudinal or long transverse direction, where fairly large specimens such as center notched panels,' are usually employed. Limitations are evident due to plate thickness, however, in the type and size of specimen that may be tested in the short transverse direction without extensions. Therefore, a specimen that is to be loaded in this direction should, for convenience, be compact. The general type of fatigue crack propagation specimens discussed and employed herein meet this requirement. These specimens are commonly called double cantilever beam specimens and lately "crackline-loaded edge-crack specimens".2 They may vary from a slope of zero (parallel-sides) to a wedge shape, the type employed herein. In general for most specimens the stress intensity KI at the tip of a crack is a function of the load, P and crack length, a. Some varieties of the wedge shaped specimen, however, give essentially a constant stress intensity KI over a considerable range of crack length.' This feature can be a valuable asset in fatigue crack propagation experiments because the stress-intensity can be controlled simply by controlling the load without regard to crack length. MATERIAL AND METHODS Material. A standard rolled (light pass reduction) and a ~reforged and rolled (heavy pass reduction) plate of 7179-T651 material were used for the evaluation. The chemistry, processing history and average tensile properties are shown in Table I. Specimen Selection and Preparation. The specimen selected for the generation of fatigue initiation or S-N data was an axial tension type and is shown in Fig. 1. Specimens were taken from mid-thickness in the longitudinal and short transverse directions from both plates. Specimens were polished with 500 grit paper in a direction parallel to the loading axis. For the fatigue crack propagation tests, the specimen shown in Fig. 2 was used. This is similar to a specimen that has been employed by Mostovoy3 for fracture toughness studies on 7075-T6 aluminum alloy. It also fortuitiously agrees quite well with the dimensions of a specimen for which Srawley and Gross2

Jan 1, 1970

By T. S. Liu, E. A. Peretti

A LITERATURE survey of the In-Sb system, which was made before a study of the binary diagram was undertaken, revealed that the intermediate phase InSb existed and that its crystal structure is face-centered cubic of the zinc blend type. In 1929 Goldschmidt' determined the lattice constant for InSb and obtained a value of 6.452A. A more recent value, 6.461A was reported by Iandelli2 in 1941. The purity of the indium used was not very high, however. The indium used by Iandelli contained 2 pct Ge. In view of this fact it was decided to remeasure the lattice parameter using high-purity indium (99.97+ pct In) and spectrographically standardized antimony of the highest purity available (99.91 pct Sb). Specimens corresponding to the InSb composition were made by carefully weighing the ingredients on an analytical balance and heating them in vacuo in a quartz tube at 700°C. As melting losses were negligible, the alloys were not analyzed chemically, but the intended analysis was taken as being accu- rate. Powder samples were prepared by either filing or crushing of the brittle alloy, followed by screening at 325 mesh. The powders were then annealed at 400°C in evacuated capsules for about an hour. X-ray photograms were taken with a Debye and a back-reflection focusing camera, using characteristic cobalt K. radiation. Lattice parameters were calculated by Cohen's³ method. The results of a Debye pattern and of a back-reflection pattern at 25°C are given in Tables I and 11. The lattice parameter at 25 °C is 6.47602A.

Jan 1, 1952

THE foundation established by the partners of the late Robert W. Hunt provides for both a medal and a prize. The latter has not been given before this year, and the first award is to C. H. Herty, Jr., for his several papers on the physical chemistry of the open-hearth process.. He is the son of the well known chemist of the same name. He was born at Athens, Ga., on Oct. G 1896, graduated from the University of North Carolina in 1918, and went into war service as foreman of U. S. Chemical Plant No. 4, at Saltsville, Va. In 1919 he went to Massachusetts Institute of Technology, where he was successively instructor, research assistant, and research associate in the School of Chemical En-gineering practice receiving his master's degree in 1921 and his Ph.D. in 1924. He continued his brilliant work on the physical chemistry of the open-hearth process, much of it being done in cooperation with the Bethlehem Steel Co. at its Lackawanna plant, and dealing with the elimination of sulfur. In 1926 he joined the staff of the U. S. Bureau of Mines, in charge of its ferrous metallur-gical work at Pittsburgh, and continues to be active in research work on the physical chemistry and metallurgy of the open-hearth process. His Institute publications are: Some Factors Affecting the Elimination of Sulfur in the Basic Open-hearth Process," "Chemical Equilibrium of Manganese, Carbon and Phosphorus in the Basic Open-hearth Process" and "Desulfurizing Action of Manganese and Iron," and he will present a paper on ?Effect of Temperature on the Solubility of Iron Oxide in Iron" at the annual meeting.

Jan 2, 1928

THE LIFE and works of Alexander Agassiz, first president of the Calumet & Hecla Mining Co., were recalled to memory when a monument bearing his statue was unveiled in Agassiz Park, at Calumet, Mich., on Sept. 26, The statue is the work of Paul W. Bartlett, of New York, and was erected by George R. Agassiz, eldest son of Alexander Agassiz. We are indebted to Rodolphe L. Agassiz, now president of the Calumet & Hecla Consolidated, for the accompanying photograph of the statue. The principal address at the unveiling ceremony was delivered by Allen F. Rees, of Houghton, who said, in part: To the world at large Alexander Agassiz was preeminently the man of science. With all his devotion to the industrial life of this district, he found the time to devote to the con-tinuation of his work of scientific research which had been the life work of his fa-ther, Louis Agassiz. Not only did he acquire eminence for his scientific at-tainments and learn-ing, but he was recognized and hon-ored throughout the civilized world for his achievements and accomplishments in the realm of original research. His work in the fields of nat-ural science, in oceanography, in marine zoology, and in other similar lines of research, has added much to our knowledge of the world's mysteries. His life was one of accomplishment also in the industrial world. The union in one life of acknowledged achievement as a scientist and the accomplishment of great results in the prac-tical industrial world is a rare thing, and its possibility has never been exemplified in greater measure than in his life and its results.

Jan 11, 1923

By Robert B. Fulton

Staurolite, an iron aluminum silicate mineral, is used industrially as a high value-in-use sand-blasting agent, as a premium grade foundry sand, and as the source of aluminum in portland cement manufacture in areas where the aluminum constituent is not economically available from shale or argillaceous limestone. It is produced by E. I. du Pont de Nemours & Co. as a coproduct in titanium mineral production in Florida where it was first recovered on a commercial scale in June 1952 from the Pleistocene sand deposit known as Trail Ridge. In the Trail Ridge deposit it comprises about one-fifth of the heavy minerals, the principal other heavy mineral constituents being ilmenite, leucoxene, rutile, zircon, kyanite, sillimanite, and tourmaline. Staurolite and the other silicate nonconductive heavy minerals are separated from the conductive titanium minerals on high-tension separators. The silicates are then separated from one another magnetically, the staurolite being relatively magnetic. In the total processing, 200 tons of raw sand yield 7 tons of heavy mineral concentrate from which about one ton of staurolite is recovered. This process yields staurolite concentrates of the characteristics given in Table 1. These concentrates are sold as three products with the size distribution of subrounded particles shown in Table 2. All three grades are used in sand-blasting for cleaning metal, for paint removal, and for cleaning buildings. Staurolite compares favorably with lower unit cost abrasives, such as boiler slags and mineral aggregates, in applications where the abrasive is used only one time and is not recovered. Because it is hard, fine, clean, and dense it can reduce blasting time, with less material reportedly being required [ ]

Jan 1, 1983

By Charles C. Upham

Not long ago a few acres of coal-land in the Connellsville region of Pennsylvania were sold at the rate of $1500 per acre. While this was doubtless a " fancy " price, affected by some consideration other than that of the simple value of the coal in the tract. it may fairly be taken as an indication of a progressive and already perceptible diminution in the economically available coal-resources of that famous region. This piesent and prospective development emphasizes the cominercial importance of producing, from the inferior coals of other districts, coke suitable for use in the blast-furnace. Many of the bituminous coals of the United States would meet this requirement if sufficiently freed, before coking, from slate and sulphur, and, so far as the removal of slate is concerned, the problem is not specially difficult. There are many forms of trough-, jig- and gravity-washers which easily and cheaply reduce the slate from as much as 20 per cent. in the coal to as little as 8 per cent. in the coke. But to reduce the sulphur from 4 per cent. in the coal (a proportion sometimes encountered) to 1 per cent. or less in the coke (the proportion demanded by the furnace-manager) is not so easy. It may be safely assumed as familiarly known to members of the Institute interested in this subject that, of the three forms in which sulphur usually exists in coal (hydrogen sulphide, calcium sulphate and pyrite), the pyrite is the one to be removed, RS far as possible, by mechanical treatment before coking. Hydrogen sulphide, being volatile, is expelled in

Jan 1, 1899

By J. D. Lubahn, R. P. Carreker, J. G. Leschen

The formation of slip bands in crystalline solids undergoing plastic deformation has recently been treated as a problem of nu-cleation and growth.' A simplified theory was developed and shown to be qualitatively consistent with certain experimental observations reported in the literature, such as the frequent failure of slip bands to traverse the entire crystal, the effect of specimen diameter and temperature on the yield stress, and the occurrence of incubation periods in the deformation of suddenly loaded crystals. The theory further predicts the appearance of transients in the rate of strain of a crystal when the applied stress is suddenly changed from one constant value to another. A transient consists of a discontinuous change in the strain rate followed by a gradual approach to a steady-state value. Such transient behavior apparently has not been recognized in previous experimental work. The preliminary experiments reported in the present paper were undertaken to confirm the existence of the predicted transients and thus to determine whether a more detailed treatment of the nucleation theory of slip might be theoretically and experimentally fruitful. Experimental Procedure A simple creep apparatus was assembled as shown in Fig I. Very little insulation was used on the furnace in order to facilitate rapid changes of temperature. Temperature was measured with a thermometer placed in the center of the furnace and was controlled by adjusting the Variac setting. The specimens were cold-headed in split-cylinder grips. The strain was measured as total elongation by a dial gauge sensitive to I/I0,000-in movement of the weight pan. Since the specimen gauge lengths were

Jan 1, 1949

By J. D. Lubahn, R. P. Carreker, J. G. Leschen

THE formation of slip bands in crystalline solids undergoing plastic deformation has recently been treated as a problem of nucleation and growth.1 A simplified theory was developed and shown to be qualitatively consistent with certain experimental observations reported in the literature, such as the frequent failure of slip bands to traverse the entire crystal, the effect of specimen diameter and temperature on the yield stress, and the occurrence of incubation periods in the deformation of suddenly loaded crystals. The theory further predicts the appearance of transients in the rate of strain of a crystal when the applied stress is suddenly changed from one constant value to another. A transient consists of a discontinuous change in the strain rate followed by a gradual approach to a steady-state value. Such transient behavior apparently has not been recognized in previous experimental work. The preliminary experiments reported in the present paper were undertaken to confirm the existence of the predicted transients and thus to determine whether a more detailed treatment of the nucleation theory of slip might be theoretically and experimentally fruitful. EXPERIMENTAL PROCEDURE A simple creep apparatus was assembled as shown in Fig 1. Very little insulation was used on the furnace in order to facilitate rapid changes of temperature. Temperature was measured with a thermometer placed in the center of the furnace and was [ ] controlled by adjusting the Variac setting. The specimens were cold-headed in split-cylinder grips. The strain was measured as total elongation by a dial gauge sensitive to 1/10,000-in. movement of the weight pan. Since the specimen gauge lengths were

Jan 1, 1948

By R. G. Stickney

INTRODUCTION The Basalt Waste Isolation Project, conducted by Rockwell Hanford Operations under contract to the U.S. Department of Energy, is assessing the feasibility of the disposal of commercial radioactive spent fuel/high-level waste in deep geologic formations in basalts beneath and within the Hanford Site in southeastern Washington State. The Basalt Waste Isolation Project mission is to identify and characterize a potential geologic repository site in basalt and to develop associated facilities and technology required for the potential permanent isolation of radioactive wastes in basalt formations. In support of the Basalt Waste Isolation Project, a laboratory testing program was conducted to evaluate the performance of the Council for Scientific and Industrial Research's "Doorstopper" biaxial strain cell. To support site characterization studies and the design of repository openings, in situ stress measurements will be conducted within the reference repository horizon (~970 m). The Doorstopper is one of the instruments being considered for use in these measurements. It was selected because it requires only a small volume of work to achieve stress relief, which is important in a closely jointed basalt. Also, test durations are short, permitting a large number of measurements to be made within a given time to increase confidence in the results. Within the reference repository horizon, temperatures are expected in the range of 40º to 60°C with a humidity of up to 100%. For this testing program, test conditions were selected to closely resemble those expected within the repository horizon. The Doorstopper test, mainly developed by Leeman (1964), is a strain rosette relief method of measuring rock stress in situ. The Doorstopper consists of a four element, 45º strain gage rosette

Jan 1, 1984

By Henry Hanson

A STUDY of the ore to be treated should be the first step in deciding on a machine for crushing or grinding. Coarse crushing is practically confined to the jaw and the gyratory crushers, the large-size jaw type being often employed in large installations as a primary crusher, followed by the gyratory type for secondary reduction. In small plants where mine-run ore is reduced in a single step, either of these machines will do the work. In intermediate and fine grinding, various machines are used. The relative merit of the different types will be governed in no small measure by the size of the installation, the character of the ore to be reduced and the final product desired for subsequent treatment. It should be possible, nevertheless, with the data available, if properly classified, to arrive at an approximation of the relative efficiency of these intermediate and fine grinders. It should be possible, also, to determine the economic grinding range of the different machines. With this in mind, it will be the object of this article to indicate, in rough outline, how the data available may be classified, rather than to attempt to discuss any particular practice and compare the results with other practices. Conclusions or generalizations can not be drawn without an exhaustive classification of data; in fact, many vital points will have to be determined by additional experiments which can be carried on most advantageously in the laboratory. The expense en-tailed in such experiments would be worth while if the data obtained should lead to definite conclusions that would aid in standardizing and simplifying grinding practice.

Jan 2, 1923

By W. L. Austin

The only essential difference among the three methods of collecting the precious metals from their low-grade ores by fusion is comprised in the nature of the vehicle in which those metals are concentrated. If lead- or coppper-ores are used, the processes are respectively lead- or copper-smelting; but when the sulphide of iron forms the matrix in which the gold and silver are collected, the process is known as pyritic smelting. The latter method presents rarely any advantages when lead- or copper-ores are available at rates which permit profitable treatment. Pyritic silver-smelting, defined by Percy as " the smelting of silver-ores, which are either free from lead or do not contain it in sufficient quantity to collect the silver, in conjunction with pyrites, in order to produce a regulus in which the silver may be concentrated," is not, as the name would imply, confined to ores of this metal alone, but embraces auriferous silver-ores, or even gold-ores simple. This method of collecting the precious metals from lowgrade dry ores, employed at Freiberg by Barthel Kohler in 1585, does not appear, up to the present, to have been introduced into this country on a working scale. And yet the possibility of getting rid in a single operation of all the earthy or siliceous gangue of a lowgrade ore and concentrating the precious metals in a matte without loss of lead (which includes gold and silver) by slagging, or volatilization, at the same time producing a highly siliceous slag, very nearly if not quite as clean as is usual in lead-smelting, certainly commends itself. In some parts of the country, owing to the com-

Jan 1, 1888

By G. R. Fitterer

BY means of the C-SiC thermocouple1, liquid metal temperatures have been found to vary much more than was previously supposed, and fortunately these variations can be directly associated with some of our former knowledge which combined to form the "art" of liquid-metal manipulation. COMMERCIAL DEVELOPMENT OF THIS DEVICE Several types of installations are operating in different plants. There are so many variables in steel plants in regard to operation, pro-cedure, arrangement of pouring platforms, ladles, etc., that most of the installations now in service in this country are "tailor-made"; never-theless, it is felt that satisfactory gains have been made inasmuch as each installation has indicated some effect of temperature upon the metal-lurgical operation, and has pointed the way toward technical and finan-cial benefits. In most cases, it has been necessary to accumulate the temperature data until sufficient metallurgical information is available for the drawing of conclusions. When mechanical difficulties have been eliminated, a distinct correlation of these factors has been obtained. The installations now in service. are used for the determination of temperatures in the basic open-hearth furnace prior to tapping and in ladles from both acid and basic furnaces prior to pouring. Some of the installations are in shops making the higher priced steels; others are in "tonnage shops." The errors of human element, smoke screens and emissivity corrections connected with the use of the optical pyrometer do not affect the C-SiC thermocouple readings. The accuracy of the C-SiC- thermocouple and the temperatures measured are discussed later in this paper.

Jan 1, 1936

By Joseph Hodgson

IN 1880, mining operations were commenced at the Copper Queen mine. The famous Queen orebody, which extended to the surface, was first quarried from a large open cut in the outcrop. The orebody was followed down to the 300 level with the Queen incline, and stoped. This, we believe, marks the introduction of square setting in the Bisbee district, which is still the method most generally used. Orebodies The orebodies in the Copper Queen mine occur in the limestones, and the main portion of the ore has been mined from the Abrigo, Martin, and. Escabrosa limestones. While the orebodies outcrop in the extreme western end of the mine, the general dip is to the east and south, at an angle of about 20°, although this dip is not by any means regular, being, in fact, very irregular locally. The ore is remarkable for its variableness in character, some of it being very soft, requiring a large amount of timber, and other portions consisting of extremely hard sulphides. The orebodies, in general, While remarkable for their continuity, are very irregular, both as to shape and size. The mine having been in operation so many years, and mining having been prosecuted over such a large area, the operations have caused the shrinkage of the overburden, and there is a constant movement, which is shown y large cracks and interstices in the surface rocks, which largely contribute to the very high deadwork or maintenance cost of the mine. It is surprising that for year 1913 it cost almost $400,000 just for the one item of repairs and deadwork in this property.

Jan 8, 1914

By C. B. Sawyer

ALTHOUGH the element beryllium was discovered as the oxide by L. N. Vauquelin in 1797, this metallic element was not produced until about 1828, and then only as an impure powder. Thereafter the greatest single contributor to our early knowledge of beryllium was Lebeau who, in about 1899, made pure beryllium crystals by electrolysis of the fluoride and then beryllium copper by the arc-furnace reduction of beryllium oxide with carbon and copper. About twenty-five years later, under the impetus of the work done by Siemens-Halske, beryllium commenced to make relatively rapid gains in industrial importance, principally when alloyed with copper. At the present half century mark, its use is firmly established in beryllium copper and in X-ray disks of pure beryllium. Its use as a moderator in nuclear processes in piles is engaging experimental attention and might become important. Other minor uses, but increasing in interest, are in nickel-base or iron-base alloys as a strengthening agent; as a protective agent against oxidation in aluminum and magnesium-base alloys; and as a ceramic in the form of the pure fused oxide. A false start was made in the wide use of beryllium oxide as a fired fluorescent powder known as beryllium-zinc silicate, a component of the inside coating of fluorescent lamps. This application, wherein general use by the public attained some importance, was abandoned a few years ago. This action was a consequence of the uneasiness which had been generated during the past decade by the sometimes fatal health hazard to the pulmonary tract now generally attributed to beryllium and some of its compounds. Numerous law-

Jan 1, 1953

By Frank F. Aplan

INTRODUCTION Coal is a solid, combustible mineral substance resulting from the degradation and alteration of vegetable matter largely in the absence of air. In this natural process of coalification, some non-combustible mineral matter is always associated with the combustible material. Coal preparation or cleaning is the process of removing this inert mineral matter from the coal. Preparation is highly successful in removing the large fragments of inert material, but is not very effective in removing the fine inorganic mineral matter intimately associated with the coal. Of the 600 million tons of coal produced in the United States every year, approximately one-half is produced by coal cleaning methods (1). The bulk of this tonnage is cleaned by coarse or intermediate gravity preparation methods (5 inches to 28 mesh, Tyler). Fine cleaning, say, -28 mesh, is done by shaking tables, washing cyclones, and flotation, or, until recently, the fines were either incorporated into the clean coal or simply discarded--not infrequently, the latter. In 1972, only 13 million tons of raw coal were cleaned by flotation in the United States, and from this 8.4 million tons of clean coal were produced (1). The percentage of clean coal produced by flotation in Europe (England, West Germany, Poland, U.S.S.R., etc.) is substantially higher. However, coal flotation has grown rapidly in the United States from only a few plants in 1950 to 31 in 1960 (2) and 66 in 1970 (1). Virtually all new preparation plants incorporate flotation into their basic flowsheet. The essential difference between ore and coal flotation is that for ores the entire tonnage is ground to flotation size, whereas for coal

Jan 1, 1976