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By G. R. Gohn, S. M. Arnold

The results of stress-corrosion tests upon several non-ferrous sheet metals are presented in this paper. The metals include tough pitch copper, five binary brasses, two leaded brasses, one copper-silicon alloy, two tin bronzes, and two nickel-silver alloys. The tests indicate that for most of these sheet metals, externally applied bending stresses below the yield point (0.01 per cent offset) do not result in stress-corrosion cracking as determined by the A.S.T.M. Tentative Method of Mercurous Nitrate Test for Copper and Copper Alloys (B 154-41T).~ Only in the cases of low brass, high brass and medium leaded brass, all cold-worked 4 B & S numbers hard, was any corrosion cracking observed below the yield strength, and in all cases such failures occurred at stresses above the proportional limit for these materials. However, when some of the same materials which showed no evidence of stress-corrosion cracking at stresses below the yield strength, are stressed beyond the yield point, stress- corrosion cracking is quite apt to occur. Since experience has shown that some of these materials, particularly the brasses, are susceptible to stress-corrosion cracking under service conditions, these tests would indicate that, besides externally applied stress and corrosion, plastic deformation in addition to that resulting from the original cold working of the material is also required to produce conditions conducive to stress- corrosion cracking.

Jan 1, 1945

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By L. L. McDaniel

Copper smelters of various kinds have operated in the Morenci district since 1872, but all have been abandoned with the exception of the present Morenci Smelter of Phelps Dodge Corporation, which was completed in 1942. During the five-year period starting in 1937, the Morenci ore body was prepared for open pit mining, pilot mill test work was carried out, and a complete reduction works, of which the Smelter is a part, was designed and erected. Actual construction work on the Morenci Smelter was started in the fall of 1940, and warming up of the units began on April 1, 1942. Charging of the reverberatory furnaces commenced on April 18, 1942, and the first anode copper was produced on April 26, 1942. The smelter was originally designed to handle the production of the Morenci Concentrator on a 25,000 ton per day program, but by the time the smelter was in operation, plans were already underway to increase the smelter capacity to handle the production of the concentrator which was being enlarged to 45,000 tons a day capacity as a war-time necessity. This extension to the smelter was completed and the new units were put in operation toward the beginning of 1944. The original smelter consisted of a smelter crushing plant, bedding plant, two direct-smelting reverberatory furnaces with two waste-heat boilers on each furnace, three converters, an anode department, a stack, and all of the usual accessory smelting equipment. The extension consisted of increasing the bedding plant from three to five beds, the reverberatory department from two to four furnaces, and from four to eight waste-heat boilers, and the converter department from three to six converters. A third converter aisle crane was added and additions were made to the flue systems and conveyor systems throughout the smelter; but no change was made in the smelter crushing plant or the anode department, and the same stack was used for all additional Smelter units. A blister casting machine was installed at that time in the south end of the converter aisle to handle excess and emergency production above the capacity of the anode department and in 1947 a converter aisle skull breaker and a lime burning plant were added as the final units for a complete plant. The choice of direct smelting over calcine smelting for the Morenci Smelter was made after careful study by members of the Western organization of Phelps Dodge Corporation and after test runs on direct smelting of Morenci concentrate had been made at the Douglas Smelter of Phelps Dodge Corporation. The Morenci furnace charge is made up of comparatively high grade concentrate with no ores of smelting grade available and with only flux, a small amount of copper precipitate and the usual amount of smelter secondaries to be smelted with the concentrate. The simplicity of direct smelting for this charge and the large amount of waste-heat steam available from direct smelting operations were factors influencing the decision to adopt direct smelting for Morenci. The design of the Morenci Smelter and the type of units selected followed best experience at the Douglas Smelter of Phelps Dodge Corporation. A description of the original smelter before operations started was given in an article in the May 1942 issue of Mining and Metallurgy. The purpose of the present article is to describe the enlarged Morenci Smelter, with a discussion of metallurgy and operating practice and to show tabulations of operating and metallurgical results obtained. Because of beginning operations during the early years of World War 11, many problems caused by labor shortage were encountered, but no major difficulties developed in starting the new plant. However, because of labor shortage, full scale Smelter production was not reached until the fall of 1946. Fig 1 shows a general plan of the Morenci Reduction Works. The arrangement of the smelter equipment is shown in Fig 2, a sectional view of the smelter is shown in Fig 3, and the smelter flow sheet is shown in Fig 4. Metallurgy The metallurgy of direct smelting, being more or less fixed by the character of the charge, is not subject to the control available in calcine smelting. Slags may be modified by the addition of suitable fluxes, but the grade of the matte is determined almost entirely by the iron:copper ratio of the concentrate. The direct smelting operation involves distributing the wet concentrate along the sidewalls and in the bath of a reverberatory furnace by means of some suitable feeding device and raising the temperature of the charge so that first the moisture is driven off, then the first-atom sulphur is eliminated, and finally the sulphide portion of the charge melts and runs into the bath, carrying with it the non-sulphide portion which has been partially fluxed to form a suitable slag. The fusion of the non-sulphide portion is completed by contact with the irony converter slag which is regularly being poured into the reverberatory furnace. The smelting rate of the charge is influenced by the mineralogi-cal composition of the sulphide portion of the concentrate and by the composition and amount of the non-sulphide portion including the fluxes added. The copper in Morenci concentrate is chiefly in the form of chalcocite, intimately associated with pyrite, and non-sulphide content is very low so that

Jan 1, 1950

By D. W. Gillmore, C. C. Wright

DEWATERING is a major problem in the preparation and utilization of fine-sized coals now being recovered in increasing amounts from colliery effluents, refuse banks, and silt ponds. Of the various methods which have, been proposed for dewatering fine coal, gravity drainage is probably the oldest and most widely used, and yet little information on this subject has been published. Consequently a laboratory investigation of gravity drainage phenomena was undertaken not only to provide basic data for predicting the drainage behavior of fine coals in cars, bunkers, basins, and silos, but also to obtain some understanding of the moisture retention properties of fine coal. Because of practical interest in the problem, a reproducible laboratory technique, which simulated plant conditions as closely as possible, was developed for the evaluation of drainage variables. The ap- paratus selected consisted of vertically-mounted lucite columns, nominally 4 to 8 ft in length and 2, 3 or 6 in. in diam; the bottom of the column was provided with a rubber stopper through which was inserted a -in. diam drainage tube; a wad of glass wool was placed on the upper part of this outlet tube to retain the coal and a plug was temporarily located at the lower end to prevent drainage during charging of the column, see Fig. 1. The most satisfactory testing procedure is as follows: 1—The fine coal, which has been thoroughly

Jan 1, 1953

R. B. BRINSMADE, Ixmiquilpan, Hgo, Mexico* (written discussion). As the discussion of Mr. Graton's paper by Messrs. Norris, Patterson, and Crane has been from the standpoint of the mineral land-value owners, a few remarks in the interest of the "under dog"-the average taxpayer and citizen-may prove timely. In my discussion of Doctor Raymond's Paper1 and of the papers of Messrs. Gibson2 and Godson, 3 have endeavored to demonstrate, from both political economy and history, that our existing system of unrestricted private ownership of land is incompatible with equality of economic opportunity and, therefore, subversive of our democratic political institutions. In Mr. Graton's paper, the same land problem is important, though camouflaged by the failure of the Federal income tax laws to distinguish between land and capital, in the requirements for valuation, or between earned and unearned incomes in the levying of taxes. I disagree with Mr. Graton's statement, on p. 2967, as follows: The immense majority of mining ventures are failures. A multitude of disappointments is relieved by occasional success. Were it not for the chance of great rewards, all the vitally necessary but mainly unremunerative work of exploration would not be undertaken. In contrast with nearly all other fields of endeavor, it is altogether unlikely that, on the whole, the gains in the successful mining ventures suffice to offset the losses in the unsuccessful.

Jan 3, 1920

By Arthur A. Brant

Let us start back as far as possible, to the beginnings of this universe, some 5 billion or more years ago. This is a time interval that can be crudely underestimated by the moon-earth tidal friction effects that have increased the length of the day by about 1/1000 second per century-from a 5 to 7 hour day to its present 24-and have moved the moon away at 5 inches per year to its present 240,000 miles. The oldest stony meteorites from interplanetary space give an age dating of some 4.6 billion years, approximating the time of formation of our earth. Further, certain stars pulsate, the period of pulsation being related to their absolute brightness. Measuring apparent brightness and comparing, gives the distance away. Thus Hubble and Shapley of Harvard, from 1925 on, were able to show that the universe reached as far as our telescopes could scan. The Doppler shift of the hydrogen red lines could only mean that the nebulae, or galaxies, were moving away at velocities increasing with their distances, by 180 km/ sec velocity increase for every million light year's distance. In short, the universe is expanding like an exploded bomb.

Jan 4, 1964

By Wido Tilmann

13.6-1. Importance. In Germany there are large tertiary brown coal reserves, most of which are located close to the surface. Therefore, brown coal has been recovered for many years by means of opencast methods. By utilizing a high degree of technology, a rational operation has been achieved, and as a result, the mining techniques have become renowned all over the world. After the war, all the dredgers and overburden spreaders, as well as the means of conveyance, had to be redeveloped for adapting this equipment to the changing conditions of the deposits. Numerous auxiliary machines were adopted from the American earth moving industry and altered according to the different requirements. As a result of such technical developments, the output per man and per shift was increased remarkably in spite of the deterioration of the geological conditions. In the whole of Germany, the total reserves of brown coal amount to 100 billion metric tons, of which 60 billion tons are situated in the Federal Republic of Germany. In the latter area, 8.5 billion tons will be recover- able by opencast means, of which 3.5 billion tons are already developed. The annual output reached 102 million tons in 1965. This total figure comes from four different mining districts: Rhineland, Helmstedt (town at the border of the Soviet occupied zone), Hesse, and Bavaria. The production figures (1965) are as follows: [ ]

Jan 1, 1968

By D. W. Gillmore, C. C. Wright

DEWATERING is a major problem in the preparation and utilization of fine-sized coals now being recovered in increasing amounts from colliery effluents, refuse banks, and silt ponds. Of the various methods which have, been proposed for dewatering fine coal, gravity drainage is probably the oldest and most widely used, and yet little information on this subject has been published. Consequently a laboratory investigation of gravity drainage phenomena was undertaken not only to provide basic data for predicting the drainage behavior of fine coals in cars, bunkers, basins, and silos, but also to obtain some understanding of the moisture retention properties of fine coal. Because of practical interest in the problem, a reproducible laboratory technique, which simulated plant conditions as closely as possible, was developed for the evaluation of drainage variables. The ap- paratus selected consisted of vertically-mounted lucite columns, nominally 4 to 8 ft in length and 2, 3 or 6 in. in diam; the bottom of the column was provided with a rubber stopper through which was inserted a -in. diam drainage tube; a wad of glass wool was placed on the upper part of this outlet tube to retain the coal and a plug was temporarily located at the lower end to prevent drainage during charging of the column, see Fig. 1. The most satisfactory testing procedure is as follows: 1—The fine coal, which has been thoroughly

Jan 1, 1953

By R. A. Grange, H. M. Stewart

Man.; steel parts may crack if quenched directly into a bath near room temperature, but not if quenched at a temperature just above the range where martensite forms and then allowed to cool slowly to room temperature. This latter procedure, which is the basis for such modern hardening techniques as "martempering" and "isothermal quenching," may entail some sacrifice in depth of hardening but very little, if any, in intensity of hardening. In planning such treatments, it is necessary to know the upper limit of the martensite range, and often desirable to know the proportion of martensite that would form on cooling to any lower temperature; furthermore, the tendency of a particular steel to crack when quenched is unquestionably associated with the temperature range of martensite formation, and consequently knowledge of this range aids in selecting the optimum composition for a given application. Certain puzzling phenomena associated with the properties of quenched and tempered steel may be at least partly clarified as the manner in which martensite forms is more fully understood. Information on the temperature range of martensite formation, particularly with respect to how much martensite results from quenching to a given temperature, is lacking for most commercial types of carbon and low-alloy steel. These considerations led us to study martensite formation in 14 carbon and low-alloy steels. The resultant data may be directly used in the following ways: (I) for selecting the lowest quenching temperature at which no martensite will form; (2) for selecting the highest quenching temperature at which virtually all marten . site will form, thereby avoiding quenching to an unnecessarily lower temperature with attendant danger of cracking; and (3) in producing a mixture of tempered martensite and bainite, which in high-carbon steels has been found to possess somewhat better ductility than tempered martensite, yet does not require the full transformation time necessary for a completely bainitic (austempered) structure. PART I. OBSERVATIONS OF MARTENSITE FORMATION IN FOURTEEN CARBON AND LOW-ALLOY STEELS Transformation of austenite to pearlite or bainite can be conveniently studied by observing the change that occurs over a period of time at constant temperature; such studies are the basis of the isothermal transformation diagram (S-curve). This isothermal technique is not applicable to martensite formation, which occurs within a time interval too small to measure, at least by any ordinary means. For all practical purposes, therefore, martensite may be regarded as forming only during cooling and not during a period of time at constant temperature. As austenite is cooled below

Jan 1, 1947

By C. S. Samis, J. C. Mitchell

Sodion activity data have been obtained for the Na-A1 system for sodium concentrations in aluminum between NN, = 25 . 10-and 300 x 10'. This concentration intert-a2 encompasses those sodium concentrations nor1)zally encountered in aluminum reduction cell operation. The sodium activity data were measured by a two-phase equilibration of sodium between a lead and an aluminam rich phase. The Na-Pb system was emplojed as a reference for the determination of the actiz7itj 0.f sodium. The activity of sodiin in aluminuwz exhibits large positi2.e deviations frorn ideality and a strong dependence, on concentration below NN, = 100 x 1K6. The inteated data indicate that the activity of aluminium is equal to its mole fraction over the range of miscibility of the Na-A1 sgstem. The data on the Na-A1 binary have been combined with experimental data on the cryolite-A1 system at 1025°C to obtain the activities of Na F and AlF, in liquid cryolite oz2er the cornposition 1-ange encountered in aluminum cell operation. The results are in agreement with activities calculated from the NaF-AlF, phase diagram. ALUMINUM is produced commercially by the electrolytic reduction of aluminum oxide dissolved in molten cryolite. There is considerable controversy over the actual mechanism of the discharge processes, but sodium potentials in the systems appear to play an important role in the reduction mechanism. The metal-salt reaction involving NaF and aluminum to produce A1F3 and sodium is responsible for the presence of sodium in the aluminum metal phase. Additional sodium may be produced from the simultaneous deposition of sodium and aluminum at the cathode as a result of a cathodic overvoltage. runert' has suggested that sodium gas at atmospheric pressure would be the result of reaction between cryolite and aluminum at 1000°C. Subsequent investigations have shown that production of sodium at atmospheric pressure does not occur. These later investigations can be conveniently separated into two categories. The first involves sodium, aluminum, and cryolite, in which the sodium activity was measured. The second includes a lead phase for sodium activity measurement. Jander and errmann,' studied the reaction: at 1090°C in alumina crucibles under an atmosphere of dry, oxygen-free nitrogen. The equilibrium constant for the reaction was expressed in terms of concentration. Their results in Fig. 1 show an increase in sodium content of the aluminum with increasing NaF/A1F3 weight ratio. Pearson and waddington3 performed similar experiments at 1000°C. Their results are included in Fig. 1. Dewing and ollinshead have measured the sodium content of aluminum as a function of the NF/AF, ratio at 1025"C, Fig. 1, and ewin' measured the sodium content of the aluminum in a system involving the solid reactants NaF, cryolite, and liquid aluminum. This reaction is different from those discussed above. Two solid phases, NaF and Na3A1F6, and a liquid aluminum phase were present. The sodium concentration in the aluminum was determined by the temperature. Dewing's measurements covered the temperature range 679" to 896°C. Omo, Matsushima, and 1t02" investigated the equilibrium 6NaF ential thermal analysis. Stokes and Frank6 determined sodium activities by

Jan 1, 1970

By J. L. Walter

Tungsten ingots with and without small amounts of aluminum, silicon, and potassium were prepared by conventional powder-metallurgy techniques and hot-rolled to 0.001-in.-thick sheet. The ingots were rolled three different ways: I) initially cross-rolled, then straight-rolled; 2) straight-rolled, and 3) straig-ht-rolled, then cross-rolled. Total reductions were 99.8 pct. Samples of the sheet were annealed, in vacuo, at temperatures from 1200" to 2200°C. Substructures and textures were determined by transmission electron microscopy and X-ray diffraction. The doping agents were found to form particles of mullite (Al,Si,O,,) which, on rolling., appear to form strings of particles aligned parallel to the rolling direction. These strings of purticles are associated with an increase in the recrystal-lizution temperature from 1400°C for undoped sheet ThE addition of small amounts of aluminum, silicon, and potassium to tungsten has a profound effect on the grain structure of drawn and recrystallized wire.'" The grains generally do not occupy the entire cross section of the wire and have lengths many times the wire diameter. The grain boundaries are, in general, nearly parallel to the wire axis and have long ''fingers" projecting one into the other as may be seen in Fig. 1, a longitudinal section of recrystallized doped wire. When unsuitable additions or no additions are made or the wire is not sufficiently reduced during draw- to 2000°C for doped sheet. The particles also limit lateral grain growth while allowing growth parallel to the rolling direction. This results in the formation by primary re crystallization of large highly elongated grains of specific preferred orientation. The undoped sheet recrystallized to small equiaxed gains. The preferred orientations of the large grains were determined from axis density plots to be (113) (691). (0i3) (362), (116) (304) types for the initially cross-rolled then straight rolled material. These orientations are similar to the orientations of the large elongated grains in doped and heavily drawn tungsten wire. The nucleation of the large grains appears to occur by recovery of certain of the polygonized regions of specific orientation, followed by strain-induced boundary migration.

Jan 1, 1968

By H. M. Stewart, R. A. Grange

Man.; steel parts may crack if quenched directly into a bath near room temperature, but not if quenched at a temperature just above the range where martensite forms and then allowed to cool slowly to room temperature. This latter procedure, which is the basis for such modern hardening techniques as "martempering" and "isothermal quenching," may entail some sacrifice in depth of hardening but very little, if any, in intensity of hardening. In planning such treatments, it is necessary to know the upper limit of the martensite range, and often desirable to know the proportion of martensite that would form on cooling to any lower temperature; furthermore, the tendency of a particular steel to crack when quenched is unquestionably associated with the temperature range of martensite formation, and consequently knowledge of this range aids in selecting the optimum composition for a given application. Certain puzzling phenomena associated with the properties of quenched and tempered steel may be at least partly clarified as the manner in which martensite forms is more fully understood. Information on the temperature range of martensite formation, particularly with respect to how much martensite results from quenching to a given temperature, is lacking for most commercial types of carbon and low-alloy steel. These considerations led us to study martensite formation in 14 carbon and low-alloy steels. The resultant data may be directly used in the following ways: (I) for selecting the lowest quenching temperature at which no martensite will form; (2) for selecting the highest quenching temperature at which virtually all marten . site will form, thereby avoiding quenching to an unnecessarily lower temperature with attendant danger of cracking; and (3) in producing a mixture of tempered martensite and bainite, which in high-carbon steels has been found to possess somewhat better ductility than tempered martensite, yet does not require the full transformation time necessary for a completely bainitic (austempered) structure. PART I. OBSERVATIONS OF MARTENSITE FORMATION IN FOURTEEN CARBON AND LOW-ALLOY STEELS Transformation of austenite to pearlite or bainite can be conveniently studied by observing the change that occurs over a period of time at constant temperature; such studies are the basis of the isothermal transformation diagram (S-curve). This isothermal technique is not applicable to martensite formation, which occurs within a time interval too small to measure, at least by any ordinary means. For all practical purposes, therefore, martensite may be regarded as forming only during cooling and not during a period of time at constant temperature. As austenite is cooled below

Jan 1, 1947

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 R. F. Jones, N. E. Truitt, M. Prats

In most published discussions and theories of in situ combustion, the combustion fronts are assumed to be vertical. However, evidence from field tests leaves no doubt that combustion fronts often advance more rapidly along the top than near the bottom of a formation as a result of difference in density between injected air and formation liquids. The approximation proposed in this paper to determine the movement of the resultant tilted combustion surfaces states that the vertical rate of movement of combustion surfaces is proportional to the horizontal oxygen flux. Where self-ignition is possible, the proposed method demands that a secondary combustion surface exist around production wells which produce some oxygen. These secondary combustion surfaces may be formed long before the primary combustion surface can advance from injection to production wells. Heat liberated near production wells at these secondary combustion surfaces can contribute to an early increase in production rate. Results indicate that significant oil recoveries cannot be obtained from the usual flood patterns (five-spots, seven-spots, etc.) without producing large volumes of unused oxygen. Ideally, to increase oxygen- consumption efficiency, well patterns should allow oil production from a first line of production wells and gas production from more distant lines of producers. However, it may be desirable to produce some gas at all wells to support (and benefit from) active secondary combustion surfaces. Results indicate that the well spacing through which combustion can be advanced is larger than that predicted by other methods. A large number of production wells may still be desirable to take quick advantage of gravity drainage. From a comparison with results at South Belridge field, California, it appears that this method adequately describes oxygen concentration and temperature histories and combustion-front shapes. However, this method does not accurately locate the most advanced point of the combustion surface. There is some field evidence to substantiate the actual presence of secondary combustion surfaces at South Belridge. Use of the proposed method appears warranted at this time when lay-over of the combustion surface can be anticipated. INTRODUCTION The assumption of vertical combustion fronts has been embodied in all previous publicationsl-6 which use the movement of combustion fronts away from injection wells to determine the temperature and fluid distributions in the reservoir. The only paper concerned with a mathematical model of the combustion process in which a nonvertical combustion front is used was written by Gottfried.11 Actually, nonvertical combustion fronts have been observed in most in situ combustion field tests for which adequate data are available.7, 8, 6 In practice, the typical vertical extent of the burned zone decreases with distance from the injection well, and this burned zone is at or near the top of the sand body. In some cases, such as at South Belridge field near Taft, Calif., the combustion surface is almost horizontal over a very large area.8 Thus, for some years an obvious and serious gap has existed between theory (vertical fronts) and practice (tilted fronts). This is indicated in Fig. 1. Tilted combustion fronts such as observed at South Belridge8 sometimes result from the natural tendency of injected gases to rise to the top of an oil sand. An initially present gas cap, even if thin, would make apparent the effect of gravity segregation from the first, particularly since formations considered for in situ combustion field operations are generally selected for their

Jan 1, 1969

By Arthur F. L. Bell

Prior to 1908 the oil production in the State of California had been almost entirely a heavy fuel oil, with a high flash point, but changed within a short period to a large percentage of refining oil with a low flash, by reason of the heavy production of light oil coming in from the Santa Maria field, the Midway field, and the increased production of the Fullerton field. This is clearly shown in the following tabulation of California's pro- The bringing in of these high percentages of light oil with their dangerous low flash and the increasing demand for the refined products necessitated the devising of some cheap and efficient means of quickly separating the lighter elements either in the field or at the main distributing points; also it became necessary to remove the excessive percentage of water which many of the light oils carried in emulsion. Union Oil Co. Port Harford Plant Probably the first important attempt to top light oil in California by other means than the ordinary stills was tried about 1908 by the Union Oil Co., at its plant at Port Harford, San Luis Obispo County, with Santa Maria oils. The process was the invention of Hubert G. Burroughs who afterward obtained a patent for the process, No. 998,837, dated July 25, 1911. The apparatus used consisted of a series of 12-in. pipes mounted in units of three pipes, one above the other, the crude oil being admitted to the upper pipe and filling it about two-thirds of its

Jan 1, 1916

By Yingchang Tsenshan Wang

UUIN L&SS 1» Paqb I. Introduction........................................................658 1. Subject and Scope...............................................658 2. Acknowledgments...............................................659 3. Review of Previous Work........................................659 Field Data...................................................659 Experiment.a1 Data............................................662 4. Plan of Present Investigation.....................................662 Method of General Procedure...................................663 I1. Experimental Work..................................................663 A. The Solution of Sphalerite..............................:........663 1. The Solvent Effect of Iron Sulphates upon Sphalerite...........663 2. The Solubility of Sphalerite in Salt Solutions of Silver, Copper and Lead...............................................668 B. Formation of Zinc Carbonates....................................674 3. The Reactions of Carbonates and Bicarbonates With Solutions of Zinc Salts............................................674 4. The Reaction of Zinc.Sulphate Solution With Limestones.......684 C. Peculiar Conditions Affectitig the Formation of Zinc Silicate (calamine) and Zinc Carbonate (smithsonite)..........................686 5a. The Effect of Carbon Dioxide Upon the Basic Silicate and the Carbonates of Zinc in Water, and Its Bearing Upon the Conditions of Theii Precipitation..............................686 5b. The Reactions of Solutions of Soluble Silicate With calamine and Smithsonite in Solution in Presence of Carbon Dioxide..689 6. The Reactions of a Mixture of Hydroxide and Silicate of Sodium with a Dilute Solution of Zinc Sulphate...................691 D. High-Temperature Methods.....................................692 7a. The Influence of Temperature on the Formation of Calamine..692 7b. Possible Effect of Heat on Calamine and Srnithsonite.........697 E. Summary of the Experimental Work..............................698 III. Petrographic Work....................................................700 1. The Occurrence of Sphalerite in Limestones or in Cherts............701 Summary of the Blende in Limestones and in Cherts.............7432 2. The Partially Altered Sphalerite and the Oxidation Products........702 a. Smithsonite with Blende....................................702 b. Calamine with Blende......................................703 c. Smithsonite and Calamine..................................703 d. Summary on the Partially Altered Sphalerite and Its Oxidation Products..............................................704 IV. Concl~~siom ........................................................-705 V. Microphotographs and Photographs with Their Explanations.............705

Jan 1, 1916

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 R. W. Raymond

The absence of a convenient glossary- of terms connected with mining and metallurgy has long been felt by the general public. It is to meet this want, not to furnish a technical manual for experts, that the following glossary has been prepared. It originated in an attempt on my part to revise for publication the manuscript of a compilation prepared from the appendix of Yale's work on Mining Titles and one or two other sources, to serve as an appendix to a new work on mining law, about to be published by Mr. H. N. Copy, of Washington, D. C. This revision soon assumed, contrary to my original intention, the proportions of a reconstruction; and with the consent of Mr. Copp, and for the purpose of receiving from my fellow-members valuable aid, I presented my still incomplete work as a paper at the Lake Superior meeting of the Institute, inviting from any quarter suggestions of new terms or better definitions to be incorporated in the glossary before its final publication. This invitation was so widely and generously responded to, that I cannot undertake to make in this place individual acknowledgments to those members of the Institute, and professional colleagues outside of it, who have favored me with assistance and advice. The labor bestowed upon this paper since its presentation at the Lake Superior meeting has considerably exceeded that of its first preparation, as

Jan 1, 1881

By Henry Fay

The erosion of guns is a complex problem which can be solved only by a detailed study of all the factors involved. In the present paper it is proposed to submit the results of observations and experiments which have extended over several years. Certain facts have been established and preliminary reports have been published.' Conclusions have been reached since then which, it is believed, are worth presenting for the purpose of promoting discussion, and with the hope that ultimately the whole truth will be known. The particular phase of the problem which will be presented here is the hardening of the inner surface of the gun tube. It has been known for a long time that after firing a large caliber gun for some time, the surface of the metal becomes hard and brittle, cracks, and wears away.2 The character of this phenomenon may be learned by an examination of some results which are typical, obtained from a detailed study of a 12-in. gun. The gun was trepanned and rings representing various sections of the gun were cut out for examination. The first ring represented the metal at the muzzle end, and the other rings were removed at approximately 120, 240, 325 and 337 in., respectively, toward the breech end. The appearance of the surface of the metal in the different sections is shown in Fig. 1. The greatest amount of wear and maximum amount of hard surface layer was found in section marked E, which is nearest the powder chamber and at the beginning of the rifing; the maximum amount of cracking in D; the maximum amount of copper deposition in C. Section B showed heat cracks on both lands and grooves, and section A showed heat cracks on the driving edge of the lands only. There was much flow of metal in both sections E and D, and there was progressively less effect noticeable toward the muzzle end. The polished cross-section of these same pieces is shown in Fig. 2. The extent of the erosion is greatest nearest the powder chamber, and diminishes toward the muzzle end.

Jan 1, 1917