Search Documents

By H. F. Moore

Jan 1, 1936

By Garth M. Crosby

Seven mines with important production records in the Coeur d'Alene lie adjacent to the Gem stocks -the Frisco (Gem), Hercules, Interstate, Rex (Sixteen to One), Success (Granite), Sunset, and Tamarack. Old records are incomplete, but avail- able data given in Table I show these mines have produced 10,550,721 tons of ore containing more than 50 million oz of silver, more than 1.5 billion lb of lead, and about 1 billion lb of zinc (approximately 1.25 million tons of lead and zinc combined). Average ore grade is 4.8 oz of silver per ton, 7.2 pct Pb, and 4.6 pct Zn. This impressive production represents about 12 pct of the District's yield. At present metal prices, with modern beneficiation, the gross value would be more than $400 million.

Jan 7, 1959

By F. R. Hensel

THE alloy under discussion is a nickel-silicon bronze and is one of many age-hardening or precipitation-hardening FIG. I.-EFFECT OF QUENCHING AND AGING TEMPERATURES ON HARDNESS OF TWO TYPICAL NICKEL-SILICON BRONZES, NO. 2351 AND No. 2362. copper-base alloys. The hardening con-stituent is a nickel-silicon compound.1 The quenching temperature necessary to pro-duce hardening by a subsequent aging treatment is a function of the composition and extends from about 700'C. to ()50'C.; the aging temperature range extends from 400° to 600°C. The work carried out by the authors of this paper was confined to cast alloys having compositions within the ranges of 1.5 to 3.00 per cent Ni; 0.4 to 0.75 per cent Si; balance, copper.

Jan 1, 1940

By E. W. Dewing

The nature of solid solutions of sodium in non-paphitic carbon at temperatures near 1000°C has been investigated by an electrolytic technique. The activity coefficient is found to vary strongly with the heat-treatment temperature of the carbon, and an analysis of the energy terms involved suggests that this is due to variations in the Fermi level. The diffusion coefficient for sodium is in the range implying that the sodium atoms (or ions) are not bound to specific sites on the carbon, in agreement with lack of evidence for specific compounds in the activity -composition curves. In the light of These results phenomena occurring in aluminum reduction cell linings are dis-cussed. It is concluded that sodium migrates by diffusion through the carbon lallice and not the pores, and that A14C9 is formed within the lining hy The excess of NaF invariably found in used linings is formed by this reaction and by air oxidation: There has recently been much interest in the swelling of carbon which occurs when it is made cathodic in an aluminum electrolysis cell, and there is general agreement that the phenomenon is due to the penetration of sodium atoms, derived from the reaction on the carbon lattice. (NaF and AlF3 are constituents of the fused-salt electrolyte.) The low-temperature lamellar compounds of potassium with graphite are well-known, and one compound between sodium and graphite has been reported.4 There is, however, virtually no information on the basic chemistry of the interaction of sodium with nongraphitic carbon at high temperatures, although Dell3 has made some fundamental studies of rates of penetration. At the time the present work was carried out these were not available. Rapoport and samoilenkol had measured rates of physical expansion in a dilatometer, using the fractional increase in length of a carbon specimen after electrolysis in a cryolite bath for 2 hr as an index of susceptibility to cathodic swelling. Work here with a similar apparatus showed that expansion (of a 1.5-in.-diam cylindrical specimen) ceased in less than 20 hr, and that the final expansion was much more reproducible than the value at some intermediate time. This suggested that an equilibrium was being set up, and that a thermodynamic approach was required. The results obtained are not as accurate as had been hoped, but they are the only ones of their type available and have enabled useful conclusions to be drawn. The experimental work was divided into two parts: an investigation of the activity-concentration diagram for sodium in carbon, and a rough determination of the diffusion rate. For the former, most of the experiments were made with petroleum coke even though this is not a form of carbon used in cell linings. It has the advantages, however, that it is a homogeneous material, that it is low in ash, and that its structure is reasonably well understood. Anthracites and coal cokes gave results of the same type as obtained with petroleum coke, but the reprodu-cibility from specimen to specimen was extremely poor. EXPERIMENTAL 1) The Activity-Composition Diagram for Sodium in Carbon. The petroleum coke was ground, pressed at 40,000 psi into plates 19 by 9 by 2 mm without added binder, and heat-treated. The sodium activities were established by an electrolytic technique. The carbon was made anodic in a cryolite bath with an aluminum cathode, and a controlled potential, E, applied so that the sodium activity at the anode was given by ln Na = 1n asa - EF/RT [21 where as, is the sodium activity generated by Reaction [I]. The relative activity is sufficient for many purposes and is simply given by The value of Na not at present known, but when the thermodynamics of the Na-A1 system have been worked out it can be derived from the equilibrium

Jan 1, 1963

By T. R. Lippard

PURE gypsum may be broken down into its constituents as follows: [ ] Standard specifications (ASTM Designation C22-25) state that a material shall not be considered gypsum if it contains less than 64.5 pct by weight of CaS042H2O. Anhydrite, the second calcium sulphate mineral found in nature, may be broken into its constituents as follows: [ ] PROPERTIES OF GYPSUM AND ANHYDRITE Pure gypsum is colorless to white but various impurities change its color to shades of gray, brown, red, or pink. It has a hardness of 2 in the Mohs scale and can be scratched with the fingernail. The specific gravity of pure gypsum is 2.3 to 2.4, so that rock gypsum, free from moisture, weighs about 145 lb per cu ft. Gypsum crystallizes in the monoclinic system in tabular crystals that exhibit good cleavage

Jan 1, 1949

By R. E. Allen

THERE was once a time when a practical oil man would appraise or buy a producing property on the basis of from $200 to $500 per barrel of average daily settled production. Curtailment-has, for the present at least, caused a virtual abandonment of the "barrel -of settled production" as a yardstick of, oil property value and in its place has come the "barrel of allowable production," which, although an artificial thing, is nevertheless now the principal factor governing the rate of ?pay out? or investment return. The potential production of a well, settled or otherwise, is of less importance in determining the present worth of a well than is the rate at which it may be produced. Speaking only of true potential for the moment, it is obvious that of two similar wells having the same potential, the one having the highest rate of allowable production would be worth proportionately more than the one of lesser allowable. The discount factor for deferred return and risk can easily vary the present values of the two wells by as much as 20 per cent when allowable rates are not greatly different.

Jan 1, 1934

By T. D. Yensen

EFFECT OF ELEMENTS ON MAGNETIC PROPERTIES OF IRON THE elements aluminum,1 arsenic,2 silicon,3 tin2 and vanadium,4,8 when used in relatively small amounts as alloying elements, seem to have a beneficial effect on the magnetic properties of iron-i. e., they increase the maximum permeability and decrease the coercive force-whereas there are definite indications that boron,5 carbon,6 nitrogen,7 oxygen,8' phosphorus,6 sulfur6 and titanium,4 as well as the metals of and adjacent to the ferro-magnetic group, nickel,9 cobalt,10 copper,11 and manganese,6 have a detrimental effect on these properties. The effects of calcium4 and magnesium4 are small, if any, while the effect of chromium 12 in small amounts is unknown and in large amounts is generally detrimental.

Jan 1, 1931

By Edwin R. Phelps, Charles W. Porterfield

BACKGROUND OF SURFACE MINING Surface mining refers to the process of removing the material (over- burden) overlying a coal seam and exposing the coal so that it can be loaded out and conveyed by trucks or other methods to the preparation plant or other facility. Surface mining has not always been important in the coal industry. In the early 1900s, almost all of the coal mined in the United States was mined by underground methods. In 1900, about 192 Mt (212 million st) of bituminous coal mere mined and less than 1% was mined by surface methods. The development and use of larger equipment has allowed surface mining to gain prominence throughout the years. Table 1 shows the total production of coal by mining method in the United States for ten-pear periods since 1920 and indicates the growth of surface-mined coal production. Surface mining will undoubtedly continue to account for a large percentage of the coal production in the US for many years because: 1) The capacity of surface mining excavators has continued to increase. Today, mining companies are using shovels (see Fig. 1) which are able to move up to 137.6 m3 (180 cu pd) in each bite and draglines (see Fig. 2) which move up to 168.2 m3 (220 cu pd) in each bucketful. 2) Increasing tonnage from the thick western coal seams is being produced by surface mining methods. 3) Surface mining has normally had lower costs and a better safety record than underground mining. 4) Surface mining normally recovers a much higher percentage of the coal reserve than can recovered by underground mining. 5) Some coal measures call best be extracted by surface mining

Jan 1, 1981

By H. M. Howe, Levy A. G.

Four aspects of flow in the plastic deformation of steel by overstrain, such as punching, wire-drawing, tensile rupture, etc, are…

Jan 1, 1915

By Robert McKay

THE practical study of corrosion requires consideration off its economic aspects. It must be based on sound scientific principles, but it should be borne in mind that probably the most important object of a corrosion test is to determine what metal in corrosion-resisting construction will produce the greatest operating profit. Thus the corrosion test must be directly connected with money values or it is not a practical corrosion test. This, perhaps unusual, connection between corrosion and economics is partially due to the fact that corrosion deals only with the time factor or duration of equipment, and it is almost the only test which deals with this feature. In these statements a departure is made from the academic definition of the word corrosion and it is done deliberately. The use of the term corrosion is unusual. It is as though we spoke of a tensile test as. a "breaking" test. We are not interested primarily in the break but in the strength of the metal. There is no English word which expresses our main interest in corrosion-testing, but let us have clearly in mind that we make corrosion tests to prevent corrosion. Our real interest is in this corrosion-prevention, not in the corrosion. Previous to 1922, it was common comment that "it is impossible to determine the course of corrosion by laboratory testing." This statement was true at that time in that the technique of laboratory testing for corrosion resistance was so vague that tests in the laboratory gave untrue results almost as often as true ones. This was so, even in the simple case of determining which one of two metals better resists a given solution. While this is no longer true the number of misleading "corrosion tests" still being ground out by laboratories is too large in the light of our present knowledge. There are two important features of corrosion testing whose neglect is a large contributing factor to misleading results: (1) The economic relations of the corrosion tests to other tests; (2) testing technique.

Jan 1, 1929

By A. T. Miesch, L. B. Riley

An error occurred in the preparation of Fig. 3, p. 249, of Basic Statistical Measures Used in Geo-chemical Investigations of Colorado Plateau Uranium Deposits. This figure shows the relation of the arithmetic mean: geometric mean ratio to the geometric deviation. Curves for the cases where n, the number of samples, is 100, 10, and 5 are too high with respect to the ordinate. The error is negligible at low geometric deviations, but becomes increasingly greater wi,th higher values of geometric deviation. Arithmetric mean values in Table I1 are not effected by the errors in Fig. 3. The authors will endeavor to supply copies of the revised curves on request; write to: A. T. Miesch, Geologist, Branch of Radioactive Materials, U.S. Geological Survey, Bldg. 25, Federal Center, Denver 25, Colo.

Jan 1, 1962

By J. Hoover Mackin

The iron ore bodies of the Iron Springs district are replacement deposits of magnetite and hematite in Jurassic limestone around the borders of three intrusions of quartz-monzonite porphyry. Production in the period 1923-1965 was about 72,000,000 tons, nearly all of which was direct-shipping ore from open pits. The intrusions are laccolithic in form; all three were emplaced at the same stratigraphic horizon, which had been a zone of decollement gliding during the Laramide orogeny. The intrusions and associated volcanic rocks are post-orogenic. There was no change in the thickness of the limestone during the metallization. Comparison of the composition of the limestone and stratigraphically equivalent replacement ore indicates that, with every million tons of iron, there were additions of 40,000 tons of silicon, 20,000 tons of magnesium, and 10,000 tons of aluminum. The ore-forming fluid was derived from the exposed porphyry. The iron was first incorporated in hydroxyl-bearing mafics, which crystallized in depth and were carried upward in the magma that formed the hypabyssal intrusions. The mafic phenocrysts decayed deuterically in the new environment, releasing iron into the interstitial fluid of a consolidating crystal mush in the interior of the intrusions. Primary tension joints, which opened in areas of late intrusive distention, served as roots through which the ore-forming fluid was drawn from the crystal mush.

Jan 1, 1968

By G. C. McCartney, S. J. Kidder

The Helen Mine, of the Algoma Steel Corporation, in the Michipicoten district, Ontario, Canada, has produced more than 6,240,290 tons of iron ore. Prior to and during World War I, 2,823,369 gross tons of brown ore were shipped, and from 1939 to the end of the 1944 season, 3,416,921 gross tons of siderite ore were produced from open-pit operations. Brief descriptions are given of the nature of the ore, the mining methods employed and the manner of treatment of the ore. Shipment of the finished product, known as Algoma sinter, is made to blast furnaces in both Canada and the United States. The iron-ore bodies—i.e., the oxide ores of the Old Helen and the siderite of the New Helen—are a part of a tabular shaped area of iron formation that extends over a length of more than 100,000 ft. and has a maximum width of 1000 ft. The ores are considered to be a product of introduction of iron-bearing material accompanied by replacement of the rocks already present. It is shown that introduction took place subsequent to the development of the major folded structures. It is believed the iron-bearing material emanated from an igneous source. Introduction The Michipicoten mining district derives its name from the Indian word1 meaning "place of bold promontories" or "region of big places," and anyone who has visited this part of Canada will agree that the district is well named. In 1897, there was active prospecting for gold in the district and the Ontario Bureau of Mines2 opened a recording office at the Hudson's Bay post at Michipicoten River. In the summer of 1897, Ben Boyer, probably searching for gold, discovered an outcrop of brown iron ore at the east end of a small lake, less than 8 miles, in an air line, northeast of the nearest point on Michipicoten Harbour. In the following year, extensive surface exploration and diamond drilling disclosed a commercial body of nonbessemer brown ore, possibly 700 ft. long and 200 ft. wide, of about the following analysis:3 Fe (natural), 57.15 per cent; S, dried at 212°F.. 0.113; SiO2, dried at 212°F., 4.00; P, dried at 212°F., 0.125. The property (Figs. 1-3) was taken over by E. V. Clergue and ultimately was transferred to the Algoma Steel Corporation. The Algoma Central Railway shortly afterward built its line from Michipicoten Harbour to the mine, together with an ore dock at the Harbour. Thus, within three years of its discovery, the mine had adequate shipping facilities and a market for its ore. Production of Brown Ore and Pyrite The first shipment of brown ore was made in July 1930. By 1939, the mine was shipping at the rate of 170,000 long tons

Jan 1, 1949

By S. J. Kidder, G. C. McCartney

The Helen Mine, of the Algoma Steel Corporation, in the Michipicoten district, Ontario, Canada, has produced more than 6,240,290 tons of iron ore. Prior to and during World War I, 2,823,369 gross tons of brown ore were shipped, and from 1939 to the end of the 1944 season, 3,416,921 gross tons of siderite ore were produced from open-pit operations. Brief descriptions are given of the nature of the ore, the mining methods employed and the manner of treatment of the ore. Shipment of the finished product, known as Algoma sinter, is made to blast furnaces in both Canada and the United States. The iron-ore bodies—i.e., the oxide ores of the Old Helen and the siderite of the New Helen—are a part of a tabular shaped area of iron formation that extends over a length of more than 100,000 ft. and has a maximum width of 1000 ft. The ores are considered to be a product of introduction of iron-bearing material accompanied by replacement of the rocks already present. It is shown that introduction took place subsequent to the development of the major folded structures. It is believed the iron-bearing material emanated from an igneous source. Introduction The Michipicoten mining district derives its name from the Indian word1 meaning "place of bold promontories" or "region of big places," and anyone who has visited this part of Canada will agree that the district is well named. In 1897, there was active prospecting for gold in the district and the Ontario Bureau of Mines2 opened a recording office at the Hudson's Bay post at Michipicoten River. In the summer of 1897, Ben Boyer, probably searching for gold, discovered an outcrop of brown iron ore at the east end of a small lake, less than 8 miles, in an air line, northeast of the nearest point on Michipicoten Harbour. In the following year, extensive surface exploration and diamond drilling disclosed a commercial body of nonbessemer brown ore, possibly 700 ft. long and 200 ft. wide, of about the following analysis:3 Fe (natural), 57.15 per cent; S, dried at 212°F.. 0.113; SiO2, dried at 212°F., 4.00; P, dried at 212°F., 0.125. The property (Figs. 1-3) was taken over by E. V. Clergue and ultimately was transferred to the Algoma Steel Corporation. The Algoma Central Railway shortly afterward built its line from Michipicoten Harbour to the mine, together with an ore dock at the Harbour. Thus, within three years of its discovery, the mine had adequate shipping facilities and a market for its ore. Production of Brown Ore and Pyrite The first shipment of brown ore was made in July 1930. By 1939, the mine was shipping at the rate of 170,000 long tons

Jan 1, 1949

By Robert J. Linney

IN the latter part of the year 1943, it was decided to experiment with diamond-drill blast holes in the Old Bed magnetite mine at the Mineville mines of the Republic Steel Corporation, in sections of the ore body where vertical dimensions were large and where development for sublevel bench stopes had already been done. Other possibilities of application for this drilling method were in the reclamation of old floors and pillars in the pitching folds of the Lower Old Bed ore body and in the reclamation of large vertical pillars in the " 21 " Bonanza-Joker ore body. The cross sections and views of the two ore bodies will show the conditions to which the diamond-drill method has been applied. The Lower Old Bed magnetite ore body is lenticular and dips approximately 30° toward the southwest and strikes generally southeast and northwest. An intensely folded structure, which pitches generally on an axis approximately at right angles to the strike, is continuous from the uppermost old workings down to the present bottom of the mine, and probably continues with depth. On successively lower levels this folded structure becomes greater in vertical dimensions and the two folds are closer together in a horizontal dimension. As shown in Fig. 1, this folded structure gives to parts of the ore body considerable vertical dimensions, which provide areas that were developed for sublevel bench stoping above a slushing drift, as shown in Fig. 2. It was in one of these previously developed sublevel bench stopes, No. 26 on the 1185-ft. level, that the diamond-drill blast-hole method was applied for mining the upper sublevel. The lower sublevel was underhand-bench-drilled to the finger raises of the scram drift below. The roof of this stoping area began to break and spall on the lower sublevel, causing the roof to become inclined. Along the hanging wall of the stope this reached an elevation about half way between the two sublevels. A smooth schistose seam in the hanging wall causes the ore to break away freely. Therefore, if bench mining had been carried on in the upper sublevel, a dangerous condition of poorly supported benches would have resulted. When the drillers were slashing across the sublevel to footwall and hanging wall to make a bench, spalling roof of the opening below would probably have carried away part of the bench, and would have left a dangerous place in which to continue underhand benching. A 6 by 8-ft. raise had been driven between the lower stope and the upper sublevel about 200 ft. from a manway. It was decided to use this raise for a starting cut and to arrange holes as shown in Fig. 3, so that a slot would be opened across the ore body between the footwall and hanging wall. This slot would then be used for the purpose of breaking into the ore; by blasting vertical rings of holes drilled at right angles to the axis of the pitching folds as nearly as possible. Layouts of this slot drilling and the rings adjacent are

Jan 1, 1945

By Robert T. Ault

The nature of fracture in unnotched tensile and notched tensile sheet and round specimens and V -notched and precracked Charpy-type sheet specimens of both wrought stress -relieved and re-crystallized molybdenum was investigated over the temperature range of -78° to 300°C. The sharp rise in fracture stress, for unnotched tensile samples, as the temperature is increased above the brittleness transition temperature (Tb), is found to be a result of an increase in flow stress due to plastic constraint and an increased strain rate which results from the onset of necking at Tb . Quasi-brittle fracture in notched tensile samples is found to occur when the tensile stress at the elastic-plastic interface, beneath the root of the notch, reaches a maximum critical value, which is independent of test temperature over the range from —78° to 25°C, but dependent on microstructure. In the temperature range between 150° and 300°C, unnotched tensile and notched tensile samples alike are found to fracture by a ductile fibrous tearing process which is discontinuous in nature, as a result of the competition between the processes of tearing, through continued plastic flow, and local work hardening. Results from the V-notched and fatigue-cracked Charpy-type impact tests demonstrate that crack initiation is the governing factor in the low-temperature (-78° to 25°C) fracture process for molybdenum. In recent years, there has been considerable interest, and a commensurate number of investigations, concerning the duc tile-to-brittle transition in refractory metals. There has not developed, however, a satisfactory explanation for the uniaxial tensile fracture-stress transition which accompanies the well-known ductility transition. This matter therefore warranted investigation. In a similar manner, the increased importance of notched tensile strength values in evaluating materials and in design criteria requires a better understanding of the factors which control the fracture behavior of notched tensile samples. In a previous investigation1 of the nature of initial yielding and fracture in notched sheet molybdenum at room temperature, it was suggested that plastic constraint was the controlling factor in governing the fracture behavior of notched samples. The final portion of this investigation was concerned with the fracture toughness of molybdenum. Of particular interest was the comparison of effective surface energies for fracture in V-notched samples with fatigue-cracked, Charpy-type samples in order to ascertain the relative importance of the initiation and propagation phases of the fracture process. MATERIALS AND TEST PROCEDURE Sheet tensile, notched tensile, and V-notched Charpy specimens were prepared from 50-mil, stress-relieved molybdenum sheet.* Half of these specimens were vacuum-annealed for 1 hr at 1200°C and 7.5 x l0-5 Torr and furnace-cooled to produce a relatively uniform grain size of 0.30 mm diameter. Round notched and unnotched tensile specimens were prepared from warm-rolled and swaged 5/8-in.-diam bar.* These specimens were vacuum-annealed for 1 hr at 1350°C and 5 x lom5 Torr and furnace-cooled to produce a grain diameter of 0.11 mm. Material analyses in ppm were: C N O H Sheet 290 10 4 1 Bar 50 30 30 2 The unnotched tensile sheet specimens had a 1/4-in. gage width and a 1-in. gage length. The unnotched tensile round specimens had a 9/32-in. gage diameter and a 1-1/4-in. gage length. The sheet and round notch tensile specimens are shown in Fig. 1. All of the notched tensile specimens tested were tandem-notched in order to study the location and mode of fracture initiation. The tandem-notched specimens were carefully machined so that the largest variation in notch depth on a single specimen was 0.001 in. Thus, when fracture occurs, the extent of plastic deformation that exists in the un-fractured notch section is that which exists just prior to fracture. All of the tandem-notched sheet specimens were electrolytic ally polished and chemically etched prior to testing. The V-notched Charpy-type specimens machined from the 50-mil sheet material had the standard dimensions of 2.125 in. long, 0.394 in. deep, and a 0.010-in. root radius. The unnotched tensile and notched tensile tests were conducted over the temperature range of -78° to 300°C, in a 10,000-lb Instron Universal Testing Machine, at a constant crosshead speed of 0.020 in. per min. The V-notched Charpy-type

Jan 1, 1964

By Eugene N. Cameron

Pegmatites of these areas have become important sources of beryl and lepidolite and have yielded cassiterite, columbite-tantalite, and other minerals. Examination of about 60 of these pegmatites leads the author to question the opinion of earlier investigators that the concentrations of beryl, lepidolite, and columbite-tantalite found have mostly formed by hydrothermal or pneumatolytic replacement of previously formed quartz-feldspar pegmatites. The author concludes, on the contrary, that present knowledge of the occurrence of minerals in zoned pegmatites can be applied to prospecting and mining of pegmatite deposits in these two areas of South West Africa. PEGMATITES occur in abundance in certain areas of South West Africa and in adjacent parts of northern Cape Province in the Union of South Africa. Some of the pegmatite deposits were prospected before World War I, and intermittent mining was done during the period between the two World Wars. The deposits of the Jooste lithium mines became an important source of lepidolite, and beryl, columbite-tantalite, wolframite, and other minerals were produced. Since World War II prospecting and mining have been carried on vigorously, with the result that South West Africa produced nearly 3000 tons of beryl during 1949 through 1953, together with more than 40,000 tons of lithium minerals, chiefly lepidolite and petalite. Small amounts of tantalite, columbite, and other minerals have also been produced during the postwar period. Most of this mineral production has come from two pegmatite areas roughly outlined in Fig. 1. One is the Karibib-Omaruru area of north central South West Africa. The second is the northeastern part of the Orange River pegmatite area, which extends into adjacent parts of Namaqualand, in northern Cape Province of the Union of South Africa. The writer visited the two pegmatite areas in July 1951. The visit was brief. but it gave an opportunity to compare some of the' pegmatites of these areas with those studied and described in detail by many investigators in the U. S., Brazil, and Canada during and since World War 11.'-l' The writer has attempted to determine, in a preliminary way, whether the concepts of internal structure and distribution of minerals in pegmatite bodies developed out of these studies are applicable to prospecting and exploration of pegmatites in South West Africa. Previous Work In an important series of papers (1929 to 1942) T. W. Gevers, H. I?. Frommurze, S. H. Haughton, G. K. Joubert, and other geologists of the Geological Survey of South Africa described and discussed the pegmatites of the two areas.*-" These works are the principal sources of information on the pegmatite mineral deposits of South West Africa. The investigations showed that in the two areas pegmatites are profusely distributed in pre-Cambrian rocks. These rocks are meta-igneous and metasedimentary units that have been folded and subsequently invaded by younger pre-Cambrian granites and by other igneous rocks. The pegmatites are considered to be genetically related to the pre-Cambrian granites. Pegmatites occur in the granite bodies themselves but are most abundant in the surrounding metamor-phic rocks, a pattern repeated in many pegmatite districts of the world. The origin of the pegmatites and the factors controlling the occurrence of mineral deposits in them were discussed at length by Gevers and Frommurze (1929)" and later by Gevers 14, 15 pp. 41-51, Ref.

Jan 1, 1956

By Donald W. Scott, Nathaniel Arbiter, A. C. Richardson

This paper describes the application of amine flotation to a specific problem—that of increasing the grade of magnetite concentrates derived from an iron ore requiring extremely fine grinding for mineral liberation. Experiments conducted on an iron ore from Wisconsin had shown that concentrates containing 60 to 62 per cent iron and 12 to 14 per cent silica could be produced by magnetic separation. The residual silica occurred as quartz and hornblende, either entrained or locked with magnetite. Most advantageous utilization of the concentrate, however, necessitated that the silica content be reduced to about 5 per cent, equivalent to an iron content of 67.5 per cent. It had been demonstrated that tabling of the magnetic concentrate removed sufficient quartz and hornblende to yield an acceptable concentrate, but disadvantages inherent in tabling dictated that flotation be investigated as a substitute. The goals set for the flotation investigation were: (I) the production of concentrates of about 68 per cent iron content from the magnetic concentrate; (2) an iron recovery of about 85 per cent of that in the feed to flotation; and (3) minimum reagent cost. Both batch and continuous flotation tests were employed; the former, to establish the procedure and the latter to verify it. Although the data were obtained on a specific ore, and with a commercial application in view, it is believed that they have a bearing on other problems where amine flotation is being used or is anticipated. Samples Two samples, designated "(minus 100-mesh magnetic concentrate" and "minus 150-mesh magnetic concentrate," were used for the batch experimental work. They differed only in the degree to which the crude ore was ground prior to magnetic separation. They were stored moist in steel drums in quantities sufficient for all the experimental work reported. This procedure had the advantage of allowing all tests to be made on common samples, but had the disadvantage that the individual flotation samples, because of the difference in storage time, were subject to various stages of surface alteration. Previous work had indicated that common samples were desirable for the batch studies, and it was decided that if these data appeared sufficiently promising the effect of surface alteration could be checked best by continuous tests. The size distributions and the assays of the samples are given in Table I. The minus 150-mesh magnetic concentrate, on which most of the flotation tests were made, assayed 60.9 per cent iron, 13.3 per cent silica, and 0.016 per cent phosphorus. The minus 325-mesh portion represented 78.I per cent of the sample by weight, and assayed 67.9 per cent iron. Thus, the insoluble, mainly silica, was concentrated

Jan 1, 1947

By Donald W. Scott, Nathaniel Arbiter, A. C. Richardson

This paper describes the application of amine flotation to a specific problem—that of increasing the grade of magnetite concentrates derived from an iron ore requiring extremely fine grinding for mineral liberation. Experiments conducted on an iron ore from Wisconsin had shown that concentrates containing 60 to 62 per cent iron and 12 to 14 per cent silica could be produced by magnetic separation. The residual silica occurred as quartz and hornblende, either entrained or locked with magnetite. Most advantageous utilization of the concentrate, however, necessitated that the silica content be reduced to about 5 per cent, equivalent to an iron content of 67.5 per cent. It had been demonstrated that tabling of the magnetic concentrate removed sufficient quartz and hornblende to yield an acceptable concentrate, but disadvantages inherent in tabling dictated that flotation be investigated as a substitute. The goals set for the flotation investigation were: (I) the production of concentrates of about 68 per cent iron content from the magnetic concentrate; (2) an iron recovery of about 85 per cent of that in the feed to flotation; and (3) minimum reagent cost. Both batch and continuous flotation tests were employed; the former, to establish the procedure and the latter to verify it. Although the data were obtained on a specific ore, and with a commercial application in view, it is believed that they have a bearing on other problems where amine flotation is being used or is anticipated. Samples Two samples, designated "(minus 100-mesh magnetic concentrate" and "minus 150-mesh magnetic concentrate," were used for the batch experimental work. They differed only in the degree to which the crude ore was ground prior to magnetic separation. They were stored moist in steel drums in quantities sufficient for all the experimental work reported. This procedure had the advantage of allowing all tests to be made on common samples, but had the disadvantage that the individual flotation samples, because of the difference in storage time, were subject to various stages of surface alteration. Previous work had indicated that common samples were desirable for the batch studies, and it was decided that if these data appeared sufficiently promising the effect of surface alteration could be checked best by continuous tests. The size distributions and the assays of the samples are given in Table I. The minus 150-mesh magnetic concentrate, on which most of the flotation tests were made, assayed 60.9 per cent iron, 13.3 per cent silica, and 0.016 per cent phosphorus. The minus 325-mesh portion represented 78.I per cent of the sample by weight, and assayed 67.9 per cent iron. Thus, the insoluble, mainly silica, was concentrated

Jan 1, 1947

By N. N. Breyer, W. R. Warke, S. Dinda

A model for the low-temperature peak, LTP, in fcc metals is presented. In high stacking fault energy metals, e.g., aluminum and nickel, it is proposed that a reversible dislocation glide and climb process might give rise to the observed high relaxation strengths and activation energies equivalent to volume self-diffusion values. In low stacking fault energy metals such as copper, gold, and silver, dislocation climb is energetically unfavorable at T < 0.5 Tm and the relaxation process is therefore mainly one of reversible glide. This model is used to explain the effect of a change in grain size and impurities on the LTP. THE low-temperature grain boundary damping peak, LTP, in spectroscopically pure aluminum,&apos; copper,17&apos; gold,2 nickel,3 and silver4 has recently been the subject of a detailed investigation by our laboratory. The main results are summarized in Table I. An empirical relationship was found between the peak relaxation strength, AE, and the width of the dissociated dislocation, do,* given by: do = µa2/24p? (Ref. 5) [I] where p is the shear modulus, y is the stacking fault energy,6-8 and a is the lattice parameter, Table 11. 4E is plotted against do in Fig. 1. Two regions are apparent: region (a)! which contains all but one of the metals, where AE decreases linearly with increasing do, i.e,, AE 1/do; and region (b), where ?E is approximately independent of do. A similar trend is observed when the ratio Qp/QL is plotted as a function of do, Fig. 2. (Qp and Ql are the activation energies for the peak and for lattice diffusion, respectively). In the past, when Qp and QL were found to be comparable, the relaxation peak was explained in terms of grain boundary sliding.10 The equivalence of Q, and Qb. the activation energy for grain boundary diffusion, prompted the suggestion that it was a consequence of reversible grain boundary migration on an atomic scale." A new model for the LTP, which is based on the experimental relationships in Figs. 1 and 2, is presented. In region (a) it is suggested that stress relaxation occurs through the reversible glide and climb of grain boundary dislocations; in region (b) only reversible glide can take place. The resultant effect on the grain boundary might be sliding or "a bowing out", but, since the total distance moved will be 50A, the atomic movements are considered of primary im- portance. It must be emphasized that this is only a qualitative model, but at the moment it appears the most likely way to explain the widely varying, but re-

Jan 1, 1969