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By Thomas Egleston

The process of lixiviating silver ores, which do or do not contain gold, by means of hyposnlphite of soda is likely to assume a very great importance in the West, the conditions being such that while it is applicable to very rich ores which do not contain lead enough to smelt, it is also equally applicable to many ores that are either too poor or too impure to be treated by any other process. Western ores are generally divided into four classes: those which contain copper enough to be smelted for copper, from which the gold and silver is extracted in the wet way, as is the practice of the Boston and Colorado works; those in which there is a large quantity of lead, which can be smelted for lead, and the gold and silver extracted from it; ores in which there is neither copper nor lead enough to allow of a process of smelting, but which can be treated in pans, these ores being " free milling " if they require no metallurgical treatment, or '(rebellious" if they have to be roasted with or without the addition of salt; and ores which do not contain enough either of lead or copper for smelting, which are poor both in silver and gold, contain large amounts of sulphur, arsenic, and antimony, and cannot be treated in many places in the West by any of these processes. Roasting with salt would convert the base metals as well as the silver into chlorides, and would give in the amalgamation a very base bullion, and the expense of the process would be SO great that the margin of profit would be very small, the reason being that for the ordinary process of pan-amalgamation, which is the only one suitable for ores containing small amounts of the base metals, and poor in silver and gold, the cost of a plant for milling is so large as not to justify the expense of treating such ores. The electrolytic processes which have been partially successful in Europe have not been tried 'here. While in the near future they will undoubtedly be used, it is hardly possible to consider them now.

Jan 1, 1884

By Strathmore R. B. Cooke

On the basis of experiments conducted on quartz using a bubble pick-up method, it was shown in an earlier paper1 that this mineral will preferentially adsorb hydrogen, calcium, or sodium ions, depending on the relative concentrations of those ions in the solution in which the quartz is immersed. For quartz particles ranging in size from 0.2 to 1 mm, it was demonstrated that the concentration of calcium in solution (assumed present as ions) necessary to completely activate quartz for flotation is given by the expression: Ca++ = [H+] X 106 + [Na+] X 10-3 In this expression, ionic concentrations are in mols per liter. It was further shown that the pick-up method is apparently more sensitive to changes in reagent concentration than the standard captive-bubble method, and that induction times are apparently much reduced. Since completing these earlier tests, a new type of cell has been constructed; this is shown in Fig 1 and 2. In Fig 1, A is a ground joint, B is a central tube reaching to within a few millimeters of the bottom of the cell, and C is a stopper which may be removed for reagent addition, and serves the further purpose of excluding carbon dioxide from the air during the test. The entire cell is constructed of Pyrex glass, and does not give the trouble experienced with the earlier cell, in which activating ions were released from the glass at high pH values. The only critical factor in the construction of the cell is the clearance between the central tube and the bottom of the cell. This clearance should be sufficiently small that the bubble can be pressed directly on the mineral grains lying on the bottom of the cell. In the pick-up tests to be described in this paper, the reagents used were all of C. P. grade, except the sodium oleate, which was Merck's "neutral powder." The quartz employed was water-clear vein quartz, sized on screens, cleaned with both hydrochloric acid and sodium hydroxide, and given a thorough final washing with distilled water. Experimental procedures were the same as described in the earlier paper. EFFECT OF SIZE OF QUARTZ ON PICK-UP To ascertain the effect of particle size on the adsorption of calcium ions, the quartz was sized from minus 14 plus 20 mesh through the intervening screen sizes to minus 270 mesh plus 400 mesh. Each size was thoroughly cleaned, and then tested in the cell at different calcium chloride and sodium hydroxide concentrations, and at a constant sodium oleate concentration of 20 mg per liter. All particles, within the size range given, exhibited complete pick-up within the curve expressed by the equation above. This presumably means, when the conditions imposed by the equation are satisfied, that this maximum is independent of particle size. However, it was found that the range through which partial particle pick-up occurred progressively broadened as particle size decreased. This is shown in Fig 3, in which curves B, C, and D show the limits at which pick-up just commences (as the pH is increased) for particles of minus 14 plus 28 mesh, minus 65 plus 100 mesh, and minus 270 plus 400 mesh size, respectively. These results indicate that for satisfactory activation, at any given pH, a lower calcium ion concentration is required for fine particles than for coarse particles. EFFECT OF HIGH ALKALINITY ON PICK-UP At calcium concentrations of between 1 and 10 mg per liter, and at high alka-linities, it was noticed that pick-up ceased as soon as calcium hydroxide commenced to precipitate. This effect was investigated at other calcium concentrations, with the same results. Solutions of calcium chloride, containing 105, 104, l03, and l02 mg of calcium per liter were made alkaline with sodium hydroxide until calcium hydroxide just started to precipitate, according to the following equation: CaCl² + NaOH -+ Ca(OH)2 + 2NaCl The beginning of precipitation was taken as that point at which either a faint opalescence appeared in the solution, or a Tyndall cone became ap-

Jan 1, 1950

By F. C. Clapp

Although many oil fields have been, and still are being, discovered in Oklahoma, the geology and structure of most of them have not become familiar to the general public because of the delay in securing government geological surveys and the reluctance of oil companies and other intercsted parties to give out their "inside information " Therefore, until official surveys are available, it behooves us to publish geological results as soon as possible. Fortunately the writer has been authorized by the Cement Field Oil Co. to publish his data on the Cement field, Caddo County, at this time. Locality and Description The Cement field is situated in the part of Oklahoma known generally until recently, as "Healdton fields," and lies 60 mi. (96 km.) northwest of the Healdton field proper. Like the Healdton field, it forms an approximate ellipse, trending northwest and southeast through the village of Cement on the St. LouisSan Francisco Railroad, on which it is reached in 21/2 hours from Oklahoma City. In its geological structure, the field constitutes an anticline over 13 mi. (21 km.) long and from 1 to 3 mi. wide; the point of greatest width being not far from its intersection by the above-named railroad. The major axis trends north 75' west from the village of Cement; but eastward appears deflected (if field interpretations are correct) to about south 45" east. About 5 mi. south lies an approximately parall el syncline, which may be conveniently called the Cyril syncline, the south and west boundaries of which may be distant many miles, but forming a closed basin south of the Cement anticfine. The positjon of the offsetting synclinal axis, which is believed to lie north and east of-Cement, has not been discovered. TOPOGKAFHY In the southwest part of the state, nearly all maps of Oklahoma show two mountain areas—Arbuckle and Wichita—which are conspicuous geological and topographic landmarks. In addition, some maps show a third, and smaller, range, named the Keechi Hills, in the vicinity of Cement. These hills also form a conspicuous feature in the landscape; but for some reason they have been neglected by geologists, and the

Jan 1, 1921

By G. L. Hundley, R. E. Mussler, D. H. Yee, F. E. Block, R. S. Olsen

An anhydrous chlorination process for the recovery of copper from chalcopyrite was investigated. Pelletized concentrate was reacted continuously with gaseous chlorine in a vertical shaft reactor at 550° and 650°C. While the copper was withdrawn with the rest of the metals as molten chlorides from the bottom, 99-x% sulfur was distilled from the top of the reactor. The molten metal chlorides were pulverized and dropped continuously into an upward flowing stream of oxygen. Over 80% of the iron chlorides were converted to oxides in a vertical free-fall reactor at 800° and 900°C while the copper chlorides passed through unoxidized. Copper was recovered in powder form by electrowinning continuously from chloride solutions using a diaphragm cell. The product sloughed easily from a titanium cathode and showed no iron contamination. Current efficiencies and power consumptions averaged 77% and 1.24 kw-hr per lb.

Jan 1, 1974

By C. L. Pillar

INTRODUCTION The location, size, character of the ore de- posit and its adjacent formations will deter- mine the mining system best used for its ex- traction. The mining system chosen will have to be engineered to suit the ore deposit and the people that will extract it. What can be done in the Philippines cannot always be accomplished in the upper Northwest Territory of Canada. Mining operations developed under a free enterprise system are planned to achieve a satisfactory return on the capital required to place them into production. Hence, the choices made in planning and developing the mining system must be dictated by sound economics. The mining system planned must favor: 1. Maximum ore extraction of the deposit with minimal grade dilution. 2. The amount of capital expenditure used to develop and equip the mine must be in harmony with the financial policy of the group yet be sufficient to provide a reliable mining system. Excessive frills and extras that are not required for efficient, safe production should not be provided. 3. Mine production costs are determined by the amount of men, material, equipment and energy used in the production process. These must be kept to minimal limits in order to insure dependable, low cost production. Extras that later are needed in the production operation can be added as required by the operator. Basically block caving is a mining system that is normally used to extract deep seated, massive, low grade deposits of copper, molybdenum and iron where stripping ratios for open cast mining are economically prohibited. The caving system is limited in its selectivity. It is a mass production type of underground ore extraction that under favorable conditions is one of the most efficient, low cost mining systems in the world today. Usually block cave mining systems are large. They are capable of producing ore at a rate of 10,000 to 100,000 tons/day from large deposits at costs that are comparable with many open cast mining systems burdened by high waste to ore ratios. Block cave mining systems differ basically in the plan of the area layout, the method of undercutting and the method of draw. The balance of the mining system is the mechanical transportation of the drawn ore to the surface, the mine supply, drainage and ventilation systems. This paper deals mainly with the methods of draw. The method of draw to be used in a block caving system dictates the development plan of the mining area, is highly influential in the method of undercutting to be employed and the supply and ventilation system to be used. There are three basic methods of draw commonly used in today's operations. They are: Gravity draw Slusher draw LHD draw The type of draw planned will have a major affect on: Capital expenditure and time required to bring a deposit into production. Productivity, efficiency and cost of the production operation. Percent recovery of the ore deposit in both tonnage and grade. Degree of production reliability Fundamentally the type-of draw used must be compatible with the physical characteristics of the deposit to be extracted and the financial position of the mining operation. The type of draw chosen must work to advantage with the size, configuration, depth, strength and fracture pattern of the deposit and its adjacent formations. It must favor the financial protection of the planned mining operation. A Gravity Method of Draw requires a finely fragmented, easy flowing product from the caving action of the deposit. To obtain the desired economic advantage the ore deposit must also have sufficient area and height of column to permit minimal development cost on a per ton basis. The gravity system re- quires the maximum amount of development openings and uses the minimum amount of mechanized equipment of all the caving systems. It is unskilled, labor-intensive and is highly sensitive to size of ore fragmen-

Jan 1, 1981

By Raymond E. Funderburk

Today, there are four alternatives available for disposal of highly toxic, non-radioactive hazardous waste: deep well injection, chemical processing, landfill and incineration. However, based upon research conducted by the author, a fifth alternative now exists. Disposal in salt. Salt, because of its chemical characteristics, is an ideal medium for retention of inorganic hazardous waste. Except for a few elemental compounds such as lithium and potassium, there is no reaction with sodium chloride. In the family of inorganic wastes, there are basically seven suites of waste that may be mixed together without any chemical reaction. The process developed by the author would be to take a suite of waste, either dewater or add moisture, add a solidification agent to the mix and slurry pipeline the mass into a cavern. Residence time before hardening can be adjusted, but the nominal time until full solidification is 48 hours. Once solidified, the mass will reach compressive strengths exceeding 3000 psi. The slurry can be introduced into an evacuated cavern pumped free of brine, or it can be placed beneath the brine and the waste will force the brine to the surface. The latter technique requires a high degree of engineering control to prevent contamination of the brine leaving the cavern-thus creating more hazardous waste and additional problems. The Environmental Protection Agency endorses the concept both at the research level and within EPA Headquarters. The technology will soon be utilized in a salt formation in Texas.

Jan 1, 1985

By William B. Plank

THE growing dependence of our vast industrial civilization (:n mineral products demands today, as never before, the highest technical skill in those who produce these product-;. That the duty of training the engineers, whose responsibility it is to produce these mineral products efficiently and in sufficient amount, falls directly on the schools of mining is beginning to be more fully recognized. The Institute, realizing this fact and that its future membership depends largely on the annual accretion of young men, is showing a keen interest in the problems of the mining schools and in the work of its Committee on Engineering Education. The Committee has, during the past several years, considered the problems connected with the training of men for the mineral industry, whether these have been problems of college training, or placement of the young engineer in the industry, or personnel training after he enters the industry. The field of interest and discussion has been as broad and varied as the mineral industry itself. Coal mining, metal mining and metallurgy have been discussed and the papers on these subjects have been outstanding contributions to the literature on mineral education.

Jan 1, 1931

By B. M. Larsen, T. E. Brower

In an earlier paper1 we discussed a simple, rapid method of taking samples of liquid steel and analyzing them for oxygen, which, though possibly not absolutely accurate (as is likewise true of all other methods), certainly yields results that are satisfactorily consistent. Accordingly, samples from hundreds of heats ill different shops were analyzed for oxygen, with results sufficient in number to be handled statistically and permit estimation of some general trends. Measurements were made in acid as well as in basic open hearths, throughout the refining period, to learn how stirring, blocking, adding ore or lime or sand, affects the oxygen content of the bath; we also followed the changes brought about by tapping, teeming, and deoxidizing. Some of these results, discussed in an earlier paper,2 bring out the dominant role of the carbon-oxygen reaction in regulating degree of oxidation both of metal and of slag, also the marked variability of concentration of dissolved oxygen [0] in excess of that which would be in equilibrium with carbon and with carbon monoxide at a partial pressure of one atmosphere. As a logical beginning on this complex problem, we take up in the present paper the variability of this excess oxygen, A[O], when other factors are as constant as they ever are in open-hearth operation; namely, toward the end of the refining period, when the bath may be regarded as in the comparatively steady state it approaches in absence of disturbing factors. This condition was chosen because by that time the significant reactions, except carbon oxidation, between slag and metal have come to substantial equilibrium; this facilitates interpretation of the complex situation and thus perinits more definite conclusions to be drawn. This excess concentration of oxygen, A[O], in the metal is needed, of course, to make the steelmaking process go at a reasonable rate; alternatively it may be regarded as a measure of the virtual pressure increment above one atmosphere, or of the degree of supersaturation, of carbon monoxide in the metal. According to our measurements, the excess concentration A[0] in liquid steel varies from about 0.009 to 0,035 per cent oxygen by weight. It is quite sensitive to change of conditions, being altered temporarily by operations such as stirring with green poles, additions of ore, lime, or sand; but as the disturbance works itself out, it resumes a level within the intermediate range characteristic of the steady state. The conditions most common in average furnace practice are covered by the following: i[O] Usual Per Cent Oxygen by Conditions Weight 1. Vigorous lime boil........... o.oog-0.01s 2. Steady state................ o.or;-0.025 3. Working of ore.............. o.023-0.035 Thus when the bath can be regarded as in a reasonably steady state—that is, when it is boiling normally, is at proper temperature with a well-shaped slag, and has been free from disturbing effects during the preceding half hour—A[0] returns from

Jan 1, 1948

By B. M. Larsen, T. E. Brower

In an earlier paper1 we discussed a simple, rapid method of taking samples of liquid steel and analyzing them for oxygen, which, though possibly not absolutely accurate (as is likewise true of all other methods), certainly yields results that are satisfactorily consistent. Accordingly, samples from hundreds of heats ill different shops were analyzed for oxygen, with results sufficient in number to be handled statistically and permit estimation of some general trends. Measurements were made in acid as well as in basic open hearths, throughout the refining period, to learn how stirring, blocking, adding ore or lime or sand, affects the oxygen content of the bath; we also followed the changes brought about by tapping, teeming, and deoxidizing. Some of these results, discussed in an earlier paper,2 bring out the dominant role of the carbon-oxygen reaction in regulating degree of oxidation both of metal and of slag, also the marked variability of concentration of dissolved oxygen [0] in excess of that which would be in equilibrium with carbon and with carbon monoxide at a partial pressure of one atmosphere. As a logical beginning on this complex problem, we take up in the present paper the variability of this excess oxygen, A[O], when other factors are as constant as they ever are in open-hearth operation; namely, toward the end of the refining period, when the bath may be regarded as in the comparatively steady state it approaches in absence of disturbing factors. This condition was chosen because by that time the significant reactions, except carbon oxidation, between slag and metal have come to substantial equilibrium; this facilitates interpretation of the complex situation and thus perinits more definite conclusions to be drawn. This excess concentration of oxygen, A[O], in the metal is needed, of course, to make the steelmaking process go at a reasonable rate; alternatively it may be regarded as a measure of the virtual pressure increment above one atmosphere, or of the degree of supersaturation, of carbon monoxide in the metal. According to our measurements, the excess concentration A[0] in liquid steel varies from about 0.009 to 0,035 per cent oxygen by weight. It is quite sensitive to change of conditions, being altered temporarily by operations such as stirring with green poles, additions of ore, lime, or sand; but as the disturbance works itself out, it resumes a level within the intermediate range characteristic of the steady state. The conditions most common in average furnace practice are covered by the following: i[O] Usual Per Cent Oxygen by Conditions Weight 1. Vigorous lime boil........... o.oog-0.01s 2. Steady state................ o.or;-0.025 3. Working of ore.............. o.023-0.035 Thus when the bath can be regarded as in a reasonably steady state—that is, when it is boiling normally, is at proper temperature with a well-shaped slag, and has been free from disturbing effects during the preceding half hour—A[0] returns from

Jan 1, 1948

By P. J. Clifford, A. B. Brown

The development of computerised mine planning has been hindered by the cost of processing the large data bases associated with mineral deposits and the inherent inconvenience of performing an interactive process on a time-shared computer. A dedicated computer is ideally suited for the purpose but reasonably priced micro/mini-computers have not had the capacity to store the volume of data or the processing capability required. Hardware development has cleared the way for the development of truly interactive mine planning software, covering the complete range of activities associated with mine planning, from geological assessment to short-term scheduling of operations, incorporating them into a single system using a common database. The integrated system, comprising software and dedicated hardware, adds previously unattainable speed, flexibility and clarity to the planning process because it has been tailored exclusively for that purpose.

Jan 1, 1983

By Charles Burrows

THE burden a man can endure depends on its magnitude and the number of tunes it is applied, as well as on many other factors. The resisting power of steel likewise is dependent on many factors. The maximum load it can support indefinitely is much less than its possible momentary burden. A light load straining the metal far within its elastic limit, if repeated indefinitely, ruptures the metal. The magnitude f the stresses that may be alternately applied and removed without rupturing the material depends on the number of repetitions of the cycle; the intervening periods of rest; the thermal, electric, and magnetic conditions, as well as the presence of secondary stresses. While we are unable to formulate the law between the magnitude f repeated stresses necessary to produce rupture and the length of time during which these alternating stress may be applied before rupture, my experience is that the time increases as the magnitude of the force decreases, according to a law which, to a first approximation, is exponential. In other words, F = Ce-t where F = magnitude f alternating stresses; = time for which stresses are continued; C = constant, which' depends on the conditions under which the operations are carried out. It is quite probable that the same type of law holds for the magnitude of a continuous force and its duration. When a rod of steel is subjected to the blows of a drilling hammer several things happen. First, one end is struck by the hammer and the other end is driven into the rock, thus causing a bodily displacement of the entire material of the steel drill. The greatest stresses involved in this bodily displacement are located at the two ends of the steel and manifest themselves by the distortion of these parts. Second, after the blow has been delivered the steel drill, due partly to its own resiliency and partly to the resiliency of the rock, experiences a vibratory motion as a whole, this motion is quite similar to the bouncing of a rubber ball thrown violently upon a smooth pavement. Third, the free vibration of the drill is determined only by its own natural period of vibration. This motion involves all the characteristics f any vibrating body emitting sound. Certain elements of the rod will experience alternations of condensation and rarefaction and, in these regions, the individual steel

Jan 6, 1921

By D. H. Davis

HAND picking was the earliest form of coal preparation, first practiced to improve the outward appearance of the coal being loaded and to remove any pieces that might appear objection- able to the buyer. It is still extensively practiced on the plus 1-in. sizes and particularly on the plus 4-in. and larger sizes. The percentage of ash reduction that can be obtained by hand picking depends entirely upon the amount of free impurity occurring in the coal. In a particularly dirty seam, where a considerable quantity of rock occurs with the coal as mined, hand picking produces sizable reductions in ash. If the coal is prepared clean at the working face hand picking is a final check; in some cases, on the smaller sizes, it is practiced only to show that an effort is being made to prepare a clean product. Because of the variety and different methods of handling coal, the early plants were not always adapted for efficient hand picking. At most mines an attempt is made to remove the visual impurities at one or more of the following places in the surface plant? (1) in the tipple during screening; (a) on picking band, belt, or shaking conveyors;*(3) during the loading of the coal as it goes into the railroad car. Whatever the physical arrangements of the plant may be, the usual practice is to run the coal over the screen, picking table or boom at such a rate that each piece of impurity is exposed to the view of the picker. Coal should be uniformly sized and spread out on the table so as not to exceed one particle in depth. The flow should be even in tonnage and the speed of travel should be from 30 to 60 ft. per minute and not exceed 80 ft. per minute. The dirtier the coal, the slower the speed required for a fixed number of pickers. Actually, picking capacity depends on the number of particles per minute that can be picked out and rejected. The percentage of impurity removed is determined by this and the average weight of piece rejected. Handling a mixture of sizes from nut to the largest lump over the same picking table or conveyor is not conducive to efficient separation by hand picking because the lump hides the smaller

Jan 1, 1943

By Satyajeet Ghosh, Edward F. Glynn

Many of the proposed techniques for calculating the reliability of rock slopes utilize Monte Carlo simulation. Monte Carlo approaches are attractive because they provide means of solving probabilistic problems with deterministic algorithms. One disadvantage shared by all simulation-based stability analyses is that they assume all the random variables are independently distributed. The paper focuses on the independence assumption and examines the effect of correlation on the probability of failure. In particular it considers the influence of possible (albeit hypothetical) correlation between joint attitudes [(8)] and joint friction angles [($)]. The study indicates that neglecting correlation can introduce significant errors into the reliability computations. With a modest positive correlation between [8] and [$] the error can lead to a 100% overestimate of the probability of failure.

Jan 1, 1982

By R. E. Reed-Hill, W. H. Hartt, W. A. Slippy

An unusual form of double kinking has been observed at the ends of {1121} twins in deformed zirconium. These kinks lie partly outside of the twin and partly inside. While they are undoubtedly closely associated with accommodating the twinning shear into the matrix, they are apparently closely related to other important aspects of twin growth. A mechanism is proposed for {1121} twin growth involving interactions between 1/3 (1120) dislocations in the bend planes and twinning dislocations. It is also shown that boundaries without long range strain fields develop at the tapered ends of the twins. The {1121} twins can, therefore, grow without forming high-energy noncoherent twin boundaries containing large numbers of closely spaced twinning dislocations. NUMEROUS investigators have indicated that slip is often closely associated with deformation twinning. An example is twin accommodation kinking,1-3 where dislocations form bend planes and permit accommodation of a twinning shear within the parent crystal. Another is the incorporation of slip dislocations into twins, as proposed by Sleeswyk and verbraak4 and Ishii and Kiho5 for iron and ß-tin, respectively, and observed in hcp metals by Price.6 In addition, it has been shown that shear discontinuities exist at the points of twin intersections in zinc,7 magnesium,8 titanium,9 ß-tin,10 and bismuth," where the continuity conditions for twin intersections, originally proposed by cahn,12 are not satisfied. A similar situation occurs for zig-zag (1121) twins in zirconium13 and ?-phase silver-aluminum alloys,14 where the twinning shears of the two component twins are not parallel. In all these latter cases additional deformation by slip is believed to accommodate the twinning shear discontinuity. The present paper is concerned with a newly observed aspect of the slip and twinning interrelation, in which two kinks, one inside the twin and the other external to the twin, are formed. These kinks, observed in conjunction with (1121) twinning in zirconium, form to accommodate the twinning shear. A model has been constructed explaining the growth and accommodation of (1121) twins in terms of these kinks. This mechanism may be significant since it is concerned with an important twin mode, which occurs not only in zirconium but also in titanium and hafnium. The practical significance of (1121) twinning has been adequately demonstrated. At 77° K the high ductility of zirconium containing a large number of grains unfavorably oriented for slip is primarily due to (1121) twinning.'5 Also, (1121) twins nucleated at 77°K may grow during further deformation at room temperature. This growth can greatly increase the transverse ductility15 as well as produce large mechanical hysteresis effects.= EXPERIMENTAL PROCEDURE Arc-melted, sponge zirconium plate previously described17 was prestrained 0.65 pct at 77° K by cold-rolling to introduce {1121} twins into the structure. Tensile specimens were cut with axes normal to the plate rolling direction. These specimens were pulled in tension to near the point of failure in an Instron machine at room temperature at a cross-head speed of 2 x10-2 in. per min. This additional deformation caused growth of the twins nucleated by prestrain. Deformed specimens were then prepared for metallo-graphic examination. This included an anodizing treatment18 that made the surfaces sensitive to polarized light, permitting small lattice rotations to be revealed. EXPERIMENTAL RESULTS Nature of the Twin-Kink Band Structure. Fig. 1 is a photomicrograph of a typical specimen after prestrain at 77° K. The structure is characterized primarily by long, narrow, parallel sided (1121) twins which norma1ly traverse the entire grain. Fig. 2 shows several (1121) twins after additional straining at room temperature. These lie near the center of the polarized light photomicrograph and are inclined upward to the right. It is evident that this latter deformation has resulted in appreciable twin growth. The dark regions extending from the ends of these twins are accommodation kinks. The difference in reflected light intensity indicates the

Jan 1, 1969

By Norman D. Fitzgerald

The purpose of this paper is to analyze the behavior of oil securities during the six-year period January 1932 through December 1937. This period was selected because of its varied character. It includes the depths of the depression, the broad and spasmodic recovery, and the recession collapse. The first part of the analysis is devoted to the correlation of a number of economic factors with an index of 13 oil stocks. The second part is a comparison of the different,ial behavior of 6 oil stocks. The relationships developed here are only a few of many that might be considered in a complete analysis. The stock market is regarded as a barometer of business confidence and as such exerts an important influence upon corporate policies, particularly upon industrial expansion. The complex and continuously changing nature of its price structure, the ebb and flow of influences and the effects of impacts require dynamic rather than static analysis. When treated in this way, many of the important fluctuations in market prices and changes in the relationships between securities appear quite logical. The market for oil securities has varied widely in recent years. The Standard Statistics index of 13 oil stocks made a bull market peak of 166.7 in September 1929. The collapse of the boom and the ensuing bear market finally carried the index down to a low of 36.4 in June 1932. From this point oil stocks made an irregular recovery; they reached a top at 139.7 in March 1937, and after maintaining this approximate level through August, declined precipitously to 93.3 in December 1937. The prices of these 13 oil stocks have gone up and down under the influence of numberless variables. These variables are of two kinds, those external to the industry, and those that are solely due to its internal condition. The external factors are political, social, and economic, both domestic and foreign. These are the influences that induce the ebb and flow of all prices in general and security prices in particular. Among the internal factors are the regulation of industry, the development of new reserves and processes, and the disequilibrium due to lack of balance between supply and demand. Some of the simpler relationships are examined in this paper.

Jan 1, 1938

By George B. Clark

Many basic engineering principles of all four phases of mining operations, namely, prospecting, exploration, development, and exploitation, can be analyzed better in terms of quantitative geology. Geological data from both field and laboratory will also complement scientific methods now being developed. THE geological data emphasized so successfully in prospecting for new deposits, that is, structural controls, strength of solutions, and type of mineralization, are basically those required for successful exploitation. In the mining of newly discovered deposits the most economical methods should be employed as early as possible to keep the overall cost per unit produced at a minimum and to permit maximum extraction of valuable minerals. A crucial question is: How can geological data be translated into useful quantitative results which will aid in achieving this end? H. E. McKinistry' has suggested that a solution may be reached in one of two ways: 1—the usual approach, use of judgment based on experience; or 2—mathematical calculations and tests on models, both subject to certain limitations. He also suggests that in addition to better use of geology more case data and theoretical data are needed on which to base sound judgment. Further research, therefore, is necessary. Perhaps in this field the emphasis should be on more specialization in mining methods and ground movement by men with thorough training in physics, engineering, geology, and underground mining. These specialists would be equipped to point out the most economical and scientific methods of exploitation. Selection of a stoping method is governed by the amount and type of support a deposit will require in the process of being mined, or by the possibility of employing the structure of the deposit to advantage in mining the ore by a caving method. In addition to these factors there are others which almost invariably influence the choice of an economical method of mining:' 1—strength of ore and wall rocks; 2—shape, horizontal area, volume, and regularity of the boundaries of the orebody, and thickness, dip and/or pitch of the deposit and individual ore shoots; 3—grade, distribution of minerals, and continuity of the ore within the boundaries of the deposit; 4—depth below surface and nature of the capping or overburden: and 5—position of the de- posit relative to surface improvements, drainage, and other mine openings. In the final analysis it is usually necessary to disregard the less important of these factors to satisfy the requirements of the more important. Because of the variation of geological conditions throughout and surrounding the deposit, no mining method will be everywhere ideally applicable to the conditions encountered in one ore deposit. The immediate problem is to interpret the above physical characteristics of deposits in terms of geological characteristics. Very few quantitative geological data are available on the factors related to a choice of mining methods. However, there are many descriptive data in mining and geological literature which collectively show how important an effect details of geology have upon all phases of mining operations. The following categories of basic mining methods were investigated to establish the geological factors that have affected their successful application: 1— open stopes with pillars; 2—sublevel stoping; 3— shrinkage stoping; 4—cut-and-fill stoping; 5— square-set mining; 6—top slicing and sublevel caving; and 7—block caving. It should be noted that the first five of these methods are listed in the order of increasing support requirements. Mines were selected as examples only where geological descriptions were complete enough to warrant their use. A study of the geological factors involved in mining operations led to a choice of the following classifications, employed in Table I: 1—structural type of orebody; 2—dimensions (geometry); 3— country rock (type); 4—faulting, folding, and fracturing; 5—alteration of ore and rock; 6—type of mineralization; and 7—geological factors determining mining method (summary). Of these factors only one yielded results that can be defined from available data in a quantitative manner, i.e., dimensions of the deposit. These are the most reliable guides that can be used in selection of suitable mining methods. They are, in general, the properties of geologic structure most difficult to evaluate by studies of models, pho-toelastic studies, and other laboratory methods, all of which are at present more limited in their applications than the geologic method. Application of geology has proved a reliable guide in other phases

Jan 1, 1955

By J. J. B. Rutherford

IT is now well known that troublesome intercrystalline corrosion may occur in austenitic stainless steels following exposure of the metal to a temperature within the range 1000° to 1500° F. (540° to 815° C.). The present paper describes a satisfactory method of estimating the extent of such corrosion, based on the measurement of the electrical resistance of a specimen before and after attack. This new method enabled us to study quantitatively the effect of several factors upon intergranular corrosion in the 18 Cr, 8 Ni type of alloy. The results show that susceptibility to this attack increases with increasing carbon content and grain size and appears to pass through a maximum for certain combinations of temperature of sensitization and period during which the alloy was maintained at that temperature; and are in complete accord with our interpretation of the mechanism of this phenomenon. The method is probably applicable to other metals or alloys subject to intercrystalline attack, such as the aluminum alloys of the duralumin type. The methods hitherto in use for estimating the degree of intergranular attack fall into three types: (1) microscopic examination,1 (2) macroscopic examination of fractures, and (3) determination of the time required for complete failure or disintegration .2 None of these lends itself very well to a measurement of the extent of corrosion. The first two are qualitative in nature and the third provides an arbitrary value largely dependent on individual judgment, since failure is a progressive phenomenon and not a sudden event. Our experience is that weight-loss determinations, although valuable in some types of corrosion investigations, bear no simple relation to intercrystalline corrosion in this type of alloy, as is to be expected since the phenomenon probably involves only the thinnest films between adjacent metallic grains.

Jan 1, 1932

By G. S. Cole

The temperature has been measured ahead of stationary solid-liquid interfaces under conditions approximating luzidirectional heat flow and therefore unidirectional solidification. Natural convection flow patterns may be deduced from the temperature distributions , temperature fluctuations, and shape of the interface. Fluid flow increases with the height and the rate of heat transfev through the interface and this is further manifested by a deviation of the interface from aflat vertical plane. The influence of fluid flow on solute inhomogeneity during alloy crystal growth can be inferred from the observed temperature distributions. A buoyancy force exists in the liquid ahead of a vertical solid-liquid interface, caused by a difference in density between cold fluid near the interface and warmer fluid in the bulk liquid. When the viscous and inertia forces in the melt exceed this buoyancy force, a flow of fluid takes place, termed natural or free thermal convection. Natural convection in purely liquid systems has been extensively studied for many years. 1"u On the other hand, fluid flow during horizontal crystal growth has oniy recently been the subject of experimental investigation."-25 The requirements for horizontal crystal growth differ from other heat flow systems. The small aspect ratio (ratio of height of cold wall to length of fluid) has never been considered. The uniform furnace gradient which supplies heat radially (and not necessarily symmetrically) differs from previous boundary conditions of uniform heat flux at the cold and hot ends. And most important of all an isothermal s/l interface is present which can adjust its shape and position to conform to heat and fluid flow. All of these boundary conditions involve complexities which cannot be readily solved analytically. Preliminary observation has demonstrated the penera1 shape of the natural convective flow pattern in transparent media.20'24'25 The flow is circulatory, directed toward the interface at the surface of the liquid, down and away from the interface at the bottom, and then up at the hot end of the melt. During crystal growth such a flow may interact with the solute boundary layer at the s/l interface to affect solute incorporation.M'1B|28'i!T Evidence has also been presented recently to show that thermal convective flow will affect the structure of ca~tin~s.~~-~~ The rate of heat transfer (conduction plus convection) in a given fluid system is a function of the temperature difference between the hot end of the liquid and the s/l interface and the height of the interface. At the lowest values of these variables* all heat is *It has been shown' that adverse temperature gradients as low as O.OOS°C oer cm are sufficient to cause convection. transferred by conduction. When the temperature difference or interface height are increased, laminar fluid flow commences and heat transport takes place by laminar convection as well as by conduction; turbulent heat transfer takes place at higher values of these variables. In the transition region between laminar and turbulent flow, boundary layer separation takes place; fluctuations in temperature are also noted and increase in amplitude and frequency as turbulence becomes dominant. In this paper, fluid flow patterns in the melt ahead of a stationary interface are deduced from observations of temperature distribution and fluctuations, heat flow rate, and interface shape. The fluid flow ahead of advancing interfaces and the effect of such flow on solute incorporation may be inferred from these measurements on stationary interfaces. Observations during enforced fluid motion will also be considered. EXPERIMENTAL PROCEDURE The metal was contained in a lava boat 10 cm in length, 1+ cm in width, and 2 cm in height, as shown in Fig. 1. A water-cooled, molybdenum block heat-sink is at one end of the boat; surrounding this assembly is a slotted stainless-steel tube noninductively wound with a nichrome heating element. Temperature was measured by 38-gage Chrome1 vs Alumel thermocouples sheathed in 0.05-cm-diam graphite-coated stainless-steel tubing, which were moved longitudinally and vertically through the melt by means of a two-dimensional motorized micrometer stage. In some experiments the thermocouple junction was exposed, but in the majority of experiments the junction was welded to the sheath tip; no significant difference

Jan 1, 1968

By Tadao Nagai, Masao Sato

Electrochemical measurements have been carried out in an auto- clave to investigate the electrochemistry of hydrogen precipitation of nickel from acetate buffered solution. After having previously obtained the rate equation for this reaction we have now made a concentrated effort to explain the rate equation electrochemically. A saturated NaC1-type Ag/AgC1 reference electrode has been improved so that it can be used in an autoclave. Polarization curves for the reaction, [H2=2H+ + 2e,] on nickel electrode were measured in solutions free from nickel ions at 185°C, and the parallel curves were obtained by changing PH2 and pH for normal and reverse reactions, respectively. Polarization curves in nickel solutions were complex, probably because the potential of hydrogen electrode is close to that of nickel and of the nickel surface altered during polarization. Then the nickel electrode potential and the reaction rate were measured simultaneously during pressure reduction at various hydrogen pressures and at various nickel concentrations. The reaction rates in terms of current density were plotted to obtain the local polarization curves for reactions Ni2+ + 2e ? Ni and H2 ?2H+ + 2e. The anodic and cathodic local polarization curves exhibit potential dependency. Therefore it is likely that the pres- sure reduction can be regarded as a coupled electrodes reaction.

Jan 1, 1973

By George H. Hanna

THE importance of the Donets coal field (the Donbas) to the national economy of the Soviet Union is well known. Great as was the significance of this tremendous deposit of coal in prewar days it is destined to become an even greater factor in the economic life of the country now that the tremendous havoc wrought by the Germans is being remedied. Since the war began, new coal fields have been extended; for example output of the Moscow and Karaganda (Kazakhstan) fields has become considerably larger while new mines are producing coal in several parts of Uzbekistan and Kirghizia. The huge West Siberian coal field, the Kuznetsk Basin, is producing more coal than ever before to supply the industries of that region and of the Urals. But so great has been the development of industry to meet war needs, especially in those districts of the Soviet Union that lie to the East of the Urals, that the coal obtained, eked out by such locally-won fuels as peat and wood, is still barely sufficient to meet industrial requirements. The main burden of the coal supply for the restored industries of the West, therefore, will fall on the Donbas.

Jan 1, 1945