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By C. Chu

A theoretical investigation has been made of the forward combustion process using a one-dimensional linear mathematical model, taking into consideration the effect of the vaporization-condensation which occurs on the leading edge of the heat wave. This work involves the solution of five coupled partial differential equations. Besides the vaporization-condensation phenomenon, these equations account for conduction, convection, combustion, heat loss, diffusion and bulk fluid flow. For the one-dimensional linear model studied, the vaporization-condensation phenomenon does not induce appreciable change in the temperature at the combustion front; and its primary effect is to create a steam plateau and to increase the length of the heated zone ahead of the combustion front. This effect becomes more pronounced at lower pressures, higher porosities or reduced gas saturations. The peak temperature and the temperature profile on the leading edge of the heat wave stabilize after a certain period. The length of the steam bank remains practically constant, although the length of the water bank increases as the heat wave advances. INTRODUCTION The existence of the vaporization-condensation phenomenon in the heat-wave process and the important role played by the phenomenon have been recognized by several investigators. Kuhn and Kochl stated that steam plateaus were frequently observed on the temperature records of their experiments. The steam plateaus were attributed primarily to the vaporization and subsequent condensation of the interstitial water existing in the oil sand. Szasz2 suggested that both lighter hydrocarbons and water are vaporized on the leading edge of the heat wave, carried forward in the gas stream, and then condensed to create banks of oil and water. Martin et al suggested that the vaporization-condensation phenomenon is one of the main mechanisms of the heat-wave process, along with thermal expansion and viscosity reduction. Wilson et al. reported the existence of a steam plateau several inches in length in their small-scale tube-run experiments. However, this important phenomenon has never been taken into consideration in the numerous theoretical analyses by various authors. 5 The purpose of this work was to study the thermal aspects of a linear heat wave, taking into consideration the vaporization-condensation on the leading edge of the wave, to determine the effect of this phenomenon on the temperature profile of the reservoir, and to investigate how this effect varies when other process variables are changed. THEORY We consider a reservoir of porous medium of cross-sectional area A, extending from x = 0 to x = L. This reservoir contains, aside from the solid matrix itself, a gas phase and a "combined liquid phase" which is a combination of two immiscible liquid phases — namely, an oil phase and a water phase. The oil present in the reservoir is assumed to consist of three fractions, a noncondensable gas, a nondistillable residuum, and a vaporizable oil fraction which may be present in both the gas phase and liquid phase. Before the heat-wave process begins, preheating has taken place and has imparted an initial temperature distribution T,(x) to the reservoir. At the start of the process, a stream of oxygen-bearing gas is introduced through the face at x = 0. This gas supports the combustion of the residual fuel and supplies the heat throughout the process. In order to develop a tractable system of equations, the model properties are postulated as follows: (1) the system is isotropic and homogeneous; (2) there is a uniform fuel content throughout the reservoir; (3) combustion takes place in a moving vertical planar region of infinitesimal thickness; (4) the temperatures of the fluids and the solid matrix are identical; (5) radiation is negligible; (6) the gas and liquid saturations are uniform throughout the reservoir and remain unchanged

Jan 1, 1965

By M. Cross

INTRODUCTION The development of refining processes involving gas injection into liquid metals has seen the evolution of a variety of designs [I]. During the last few years or so the top, bottom and side blown processes have, essentially, been combined to produce a series of mixed gas blowing schemes (eg. the LBE [2,3] and LD-OB processes [4,5]). Although, a variety of mixed gas blowing processes have been carefully engineered and recorded operational experience now attests to their effectiveness, optimising their design and operational performance is another matter. Detailed process analysis of such operations is impeded by the complex nature of the physical and chemical processes involved: . fluid flow behaviour of the liquid metal and injected gases, . chemical reactions in the vessel, . mass and heat transfer within the fluid phases in the vessel, . influence of the geometry (eg. location of the tuyeres, etc.), . monitoring the operational performance. By and large, design and operating parameters evolve from a combination of laboratory scale experiments, pilot plant work and, of course, practical engineering experience. Conventionally, the gas-liquid system used to model the industrial process is air-water, for a number of valid reasons including – . visual demonstration plus geometric and (to some extent) dynamic similarity, . the fact that the system demonstrates, at least, qualitatively many of the complex features of gas injection - interaction between the gas and the liquid - the nature of the gas flow behaviour - qualitative indications of mass transfer rates Important experimental contributions to the understanding of gas liquid systems include the work of Brimacombe [6], Wraith [7], T.Robertson [8], D.Robertson [9], Guthrie [10] and Szekely [11] (these references provide starting points to identify the major published work in this area). Although air-water systems have been instrumental in providing a great deal of qualitative insight into the fluid flow behaviour of the processes, reliable quantitative predictions are more problematic. A number of workers (notably Brimacombe and colleagues [6, 12])have demonstrated that different gas-liquid systems can lead to wide variations in behaviour. Other limitations of physical modelling include the assumptions of isothermal flow, scale- up of turbulent flow behaviour and the inability to directly include the effects of chemical reactions. The above limitations in physical modelling, therefore, yield a natural role for mathematical modelling provided that it can be made comprehensive enough whilst remaining cost-effective. The major physical processes involved in gas injection systems may be summarised as . behaviour of the gas envelope structure in moving through the liquid metal . the global movement of the liquid metal . momentum interaction between the liquid metal and the injected gas . heat transfer and mass exchange via the chemical react ions. The behaviour of gas envelopes forming at a tuyere and then moving through the liquid has been investigated thoroughly by a large number of workers. Such models are generally based upon simple physic- al concepts derived from visual observation. An early model by Davidson and Schuler (DS), based upon a force balance between the liquid inertia and a combination of the bouyancy and momentum forces on the envelope, has proved to be a useful tool in predicting bubble volumes at detachment [13]. The DS model has been successfully modified in a number of ways [14-17] to cope with heat transfer and reactive transfer rates, etc. Furthermore, the basic ideas summarised above have also been confirmed on industrial converters by both Guthrie [I8] and Brimacombe [19]. Models have also been developed to predict the behaviour of the "chemical" component of some gas-liquid systems. However, one thing is clear; all gas-liquid reactive systems are governed by the degree to which effective mixing occurs. As such, the processes are dominated by the behaviour of the heat and mass exchange of the two phase fluid flow system. It is the pattern of movement of the turbulent liquid metal plus its interaction with the injected gas which must be predicted reliably, if sufficiently comprehensive comp-

Jan 1, 1984

By J. L. Bruce

For a number of years, from an economic standpoint, the ventilation of the Colorado mine of the Davis-Daly Copper Co. has been a difficult problem. The development through the main hoisting shaft was pursued for a number of years and to a depth of about 2500 ft.( 762 m.)before large bodies of good commercial ore were opened up. This lack of early success in exploration restricted expenditures for ventilation, with the result that the present orebodies were opened up at a depth of 2500 ft., and in country rock f relatively high temperature, without adequate means f supplying the development faces and the stopes with cool or fresh air of low humidity and suitable velocity. During the progress of development, two or three connections were made at higher levels with the workings of adjoining operations, but these were of a makeshift and unsatisfactory character and afforded no means for carrying downcast air from the surface, in anything approaching a direct course, to the working faces. During the latter part of 1917, arrangements were made with. the Anaconda Copper Mining Co., whereby, during the year 1918 and the early part f the year 1919, a crosscut was driven eastwards on the 2500-ft. level of the Colorado shaft and connection was made by means f an up-raise of approximately 200 ft. to a crosscut driven westwards from the 2800-ft. level of the Belmont mine, this crosscut at the top of the raise being driven at Davis-Daly expense by the Anaconda Copper Mining Co. This connection, in conjunction with a No. 11 Sirocco fan installed near the Colorado shaft, provided the Colorado mine with a maximum of approximately 30,000 cu. ft. per min. of relatively fresh air, most of which came in a fairly direct route from the downcast Belmont shaft through the No. 11 fan to the 2500-ft. level stopes. After passing through these stopes and the stopes of overlying levels, it passed through the main crosscuts on higher levels back to the Colorado shaft, which provided the upcast for this current of air. The improvement

Jan 9, 1921

By W. A. Cole, E. L. Derby, J. F. Wolff

Exploration of Lake Superior iron ores is done principally by drilling. The soft iron ores are churn drilled and the harder ores are diamond drilled. The churn drill used is the Mesabi type and its operation should be described as "water jetting." The drill rods are washed down by the combination action of the water jetting from the bit at the lower end of the rods and the churning motion given the line of rods. The water, which is forced down through the rods, comes up between the wash rods and the casing, bringing the choppings to the surface. It is caught in three or more barrels, each of which has a stoppered hole 8-10 in. from the bottom. These barrels are successively filled, allowed to settle, and drained to the hole, until the desired depth is drilled. The sediment from each barrel is transferred to a settling tub in which it is allowed to settle; the water is then drawn off and the wet sample is collected and dried. The number of barrels used depends on the character of the ore. For example, when drilling sandy ores, all the wash water is caught and settled for the proper length of time without refilling any of the barrels. Thus, the sediment held below the stoppered hole in any one barrel is not disturbed. The casing is kept as near the bottom of the hole as practicable to prevent the caving of the hole from jeopardizing the sample. While the casing is being driven down and the hole is being washed clean after casing, the return water is not run into the sample barrels. In exploratory work in the harder formations, where the diamond drill is used, the sample comprises both core and cuttings. The care given in handling and preserving the core depends on the importance that may be attached to certain characteristic strata in different formations. The following instructions for the care of core samples are issued by the Ogle-bay Norton Co. for diamond drill work done for them on the Gogebic Range of Michigan:

Jan 1, 1925

By C. E. Mendenhall

THE word temperature has both a colloquial and a technical use. For everyday purposes of abusing the weather man, no very exact definition is necessary, but for the purpose of giving a simple summary of the physical laws that form the basis of practical pyrometry, something more precise is required. Beginning, therefore, with the common concept of "hotness" and "coldness," we must agree on a method of measuring differences in "hotness," on the unit to be used, and on the point from which measurements are to be taken. We shall then have a definite "scale of temperature," which can be used in all methods of pyrometry. But, as in many similar cases, it is much easier to define or describe something than it is to make practical application and use of the definition; so that much of our attention will be taken up with practical methods of realizing or applying the scale agreed upon. It was early observed that changes in temperature produced large and easily measurable changes in gases, which may be most simply separated into changes in volume (expansion and contraction) under conditions of constant pressure, and changes in pressure under conditions of constant volume. These changes are much the same in magnitude for the common gases oxygen, nitrogen, and hydrogen and also for the rarer helium and argon. This relative uniformity in behavior led to the suggestion of "the gas thermometer" and "the gas scale" as the basis for all temperature measurements. However, as methods were refined, differences appeared between different gases and different ways of using gases, so that Lord Kelvin introduced his more fundamental notion of the "absolute thermodynamic scale" of temperature, which he defined as follows: Given two bodies, say two tanks of water, at different temperatures to determine these temperatures on the "absolute thermodynamic," or Kelvin, scale, operate a thermal engine between these two temperatures, letting it take in heat from the hot body and give out heat to the cold, which therefore corresponds to the boiler and the condenser of a steam engine. The engine we may imagine as a cylinder and piston inclosing a gas and operating with the well-known Carnot or isothermal-

Jan 9, 1919

By Edward F. Fitzhugh

Quantitative appraisal of the chances of finding various tonnages and grades of ore clarifies any exploratory development proposal. Ways are discussed of appraising chances in conformance with probability principles, and graphs are suggested to show tonnages, grades and the corresponding judged expectancies. Introduction When exploratory development of a mineral deposit is contemplated, three interrelated questions normally must be answered before a sound decision can be reached. How much will the program cost? How much will the anticipated ore add to the immediate value of the mine? And what is the chance of developing this ore? To justify development, the deposit should present a favorable disparity between the three factors, cost, chance and return. If, for example, there is a 50-50 chance of success, the ore that is sought should have an immediate worth (after deducting all subsequent charges that it must bear) of more than twice the cost of the development. Likewise, if the chance of finding the ore is but one in five, that ore should enhance the immediate value of the mine by more than five times the cost of developing it- With sound economic and operating assumptions, the cost of exploratory development and a present value of anticipated ore may be estimated within reasonable limits. Less general attention has been given to appraising ore chances in comparable terms. The exploratory experience of the mining industry is broad enough that knowledge, vision and judgment may be combined to yield more tangible evaluations of Ore expectancy than are now customary' A way of making such appraisals is suggested herein. The mathematics of probabilities cannot be applied directly to the search for ore. By coupling applicable principles of probability, however, with pertinent experience gained in mineral exploration, an estimate often can be made of the chance of finding a certain tonnage and grade of Ore in a specific location. Such estimated, or judged, expectancies for the several blocks of ground in a deposit may be combined in a simple chart to show the relationship of the various exploratory possibilities and the potential magnitude of the deposit' An engineer or geologist may find that the quantitative thinking required for an appraisal of this type is an aid in clarifying and coordinating his views and recommendations. For a person wishing to understand these views, an ore-expectancy

Jan 1, 1949

By Edward F. Fitzhugh

Quantitative appraisal of the chances of finding various tonnages and grades of ore clarifies any exploratory development proposal. Ways are discussed of appraising chances in conformance with probability principles, and graphs are suggested to show tonnages, grades and the corresponding judged expectancies. Introduction When exploratory development of a mineral deposit is contemplated, three interrelated questions normally must be answered before a sound decision can be reached. How much will the program cost? How much will the anticipated ore add to the immediate value of the mine? And what is the chance of developing this ore? To justify development, the deposit should present a favorable disparity between the three factors, cost, chance and return. If, for example, there is a 50-50 chance of success, the ore that is sought should have an immediate worth (after deducting all subsequent charges that it must bear) of more than twice the cost of the development. Likewise, if the chance of finding the ore is but one in five, that ore should enhance the immediate value of the mine by more than five times the cost of developing it- With sound economic and operating assumptions, the cost of exploratory development and a present value of anticipated ore may be estimated within reasonable limits. Less general attention has been given to appraising ore chances in comparable terms. The exploratory experience of the mining industry is broad enough that knowledge, vision and judgment may be combined to yield more tangible evaluations of Ore expectancy than are now customary' A way of making such appraisals is suggested herein. The mathematics of probabilities cannot be applied directly to the search for ore. By coupling applicable principles of probability, however, with pertinent experience gained in mineral exploration, an estimate often can be made of the chance of finding a certain tonnage and grade of Ore in a specific location. Such estimated, or judged, expectancies for the several blocks of ground in a deposit may be combined in a simple chart to show the relationship of the various exploratory possibilities and the potential magnitude of the deposit' An engineer or geologist may find that the quantitative thinking required for an appraisal of this type is an aid in clarifying and coordinating his views and recommendations. For a person wishing to understand these views, an ore-expectancy

Jan 1, 1949

By C. E. Sims

IN an earlier paper by Sims, Moore, and Williams: data were given to show that hydrogen contents of 0.3 relative volume (0.00033 wt pct), existing at the center of 4-in.-square cast-steel coupons, could be removed by aging at 400°F for 125 hr. At the end of that time, complete restoration of normal ductility for the steel was obtained. Replicate specimens of the same steel were aged at room temperature up to 555 days or .18 1/2 months. At the end of that time, the hydrogen content at the center had fallen to 0.13 relative volume or 0.00015 wt pet, but there was no perceptible change in the ductility. On the basis of experience with accelerated aging of 1-in. and 4-in. coupons and room temperature aging of 1-in. coupons, it was predicted that, some time after two years and when the hydrogen content has been reduced to about 0.10 relative volume, there would be some recovery of ductility. One set of three bars remained, and these were held for 1310 days, or more than three years. Then tensile test bars were cut from the center, and these bars were sampled for hydrogen analysis at the time they were broken. The sample specimens were cylinders % in. in diam x 1 1/4 in. in length. The data showed that in this time the average hydrogen content had fallen to 0.096 relative volume or 0.00011 wt pct, and the ductility had increased sharply. Elongation had risen from 10 to 17.3 pct, while reduction of area increased from about 16 to 27.5 pct. On the basis of these data, fig. 1 was drawn to replace fig. 15 of the earlier paper. Extrapolation of the curves of fig. 1 indicates that it would require a period of about 2200 days or approximately six years at room tem- perature to reduce the hydrogen to the minimum value and to recover full ductility in the center of the coupons. The aging process apparently involves the movement of hydrogen from the interior to the surface of the bar and evaporation at the surface. This flux of gas presumably follows some diffusion mechanism, and an effort was made to use the aging data, obtained at room temperature and at 400°F, to calcu-late coefficients of diffusion, in the hope that they might indicate the nature of the diffusion. The conclusion was reached. however. that the data were inadequate for such a purpose.

Jan 1, 1951

By Gerald Sherman

IN the southern field of the anthracite region of Pennsylvania, many of the coal seams dip at angles of 40° to 50° from the horizontal. For transportation purposes, "tunnels" are driven across the coal measures and, occasionally, "rock gangways," beside the seams and parallel with them. The use of gangways driven in country rock appears to be increas-ing, as they avoid the maintenance' of long gangways in coal, which eventually are driven to the colliery boundaries, and permit mining from a number of intermediate blocks without interference. The rock consists of slates or sandstones of moderate hardness, which may range into very hard, highly siliceous, fine-grained conglomerates. In this distinctly bedded formation,' the direction of the heading, whether parallel with or across the structure, makes it necessary to distribute the drill holes in a different pattern to break the ground, but does not change loading conditions. In 1932, practically all the mucking was done by hand, and it was the established custom to use four men on two drills for drilling and breaking the ground, and five for loading it, out. A driver and mule was needed in addition, to assemble the rock cars at a siding or branching track in the rear, pending their final disposal. This might occupy the driver's whole time, or if the length of haul were short, he might be able to do other useful work. It was thought possible in 1932 to reduce the cost of driving such headings in rock by mechanical loading, and to investigate this subject a scraper loader was used in a section of tunnel then being driven. Opera-tions while driving 195 ft. of this tunnel using a scraper loader were care-fully observed. Time studies were made of the loading, and records made of the routine of breaking ground, but not in such detail. Tunnels or gangways in the colliery are 7 ft. high above the rails, by 10 ft. wide, which requires a minimum of 8 ft. in height, broken to allow for rails and ties. This cross section is of sufficient size to permit the use of any one of several types of power loaders, but for this demonstration if was found most convenient to use a scraper loader.

Jan 1, 1936

By Paul Tyler

FULLY a decade ago, demand for ground mica began to exceed supplies of scrap mica from manufacturing operations and of waste block from feldspar and sheet mica mining in the United States, with the result that in recent years the market has absorbed larger and larger quantities of mica recovered as a by-product from clay washing,1 as a joint product with kyanite,2 and by grinding schists. So far, only schists that contain a large percentage of mica have been utilized for making ground mica, but already there has been a good deal of wishful thinking with regard to recovering mica as an aid to the prospective working of low-grade gold ores and ores of certain base metals. With wet-ground mica selling at $60 to $100 a ton and even dry-ground mica worth over $20 a ton, many ores that could not be considered as otherwise workable under present conditions will show commercial possibilities if they can be made to yield even a relatively small percentage of salable mica. An interesting paper might be written on the coming role of industrial minerals as joint products of metal mines, It is largely a matter of mar-kets. Many years ago I watched the mess left behind the gold dredges in the mountains of Colorado and thought of the Denver sand and gravel operators who prided themselves upon their canniness in providing devices for catching gold merely as an incident to their regular business of selling what we miners classed as "tailings." Nowadays almost every milling or mining operation near settled communities finds use for at least part of its waste for road material or concrete aggregate and several eastern iron mines have gone into the crushed-stone business. Promoters now mention fertilizer "filler" as potentially salable at all the way from $1 to $10 a ton, or whatever is needed to cover costs of mining and treating the "ore" and leave the metal values as "clear profit." In emulating the pork packers in their efforts to utilize "everything but the squeal,"

Jan 1, 1938

By Henry Howe

W. E. RUDER, Schenectady, N. Y.-In 1913 I presented a paper before this society on Grain Growth, and at that time it seemed to me that the only explanation for grain growth was that of critical strain. I do not think I published the photographs at the time, but I made some experiments on the effect of stretching sheets of metal. This allowed for fairly accurate determinations of the amount of cold deformation, At the time I worked-with silicon steel because one could see the change in grain growth with the naked eye. I stretched about 10-samples from 0.5 per cent. to 10 per cent., and, after heating to 1,050° C., I found that the maximum grain growth occurred at about 2 ½ per cent. cold deformation. This I designated "the critical strain" for maximum grain growth. About a year ago I talked with Prof. Jeffries, who explained to me his ideas of grain growth. It occurred to me that it would be a very easy thing to check the strain hypothesis against the strain and temperature hypothesis, if we took a certain number of these strips, strained exactly alike and heated to different temperatures; in that case the maximum grain growth ought to appear at a different percentage of deformation for each temperature. These experiments showed that there was, as Mr. Jeffries shows in his present paper, a decrease in germinative temperature with increased cold deformation. The samples were stretched from ½ to 8 per cent. and heated for 10 hr. in hydrogen. Samples were introduced into a furnace previously heated to temperature. A maximum grain growth was obtained at 5 per cent. deformation, when heated to 750° C.; at 2 ½ per cent. when heated to 800' C.; at 1 1/4 per cent., if heated at 950°C.; and at 5/8 per cent.. when heated at 1,100° C. This shows very clearly that the germinative temperature decreases with the increase in cold deformation. The growth was very much more marked, however, at the high temperature, i.e., the grain grew more rapidly and attained a much larger size than at lower temperature. In considering the photographs (Fig. A) those along the lateral edges should be ignored because.

Jan 4, 1917

By Lucien Eaton

The use of filling in mines is less common in the United States than it is in Europe, where in some places it is required by law. In most cases the filling is placed by hand, and the material used for this purpose is the waste from development and that sorted from the ore in the stopes. When these sources provide an insufficient supply, additional rock must be quarried on surface or mined underground, or some other material must be used. In different parts of the world sand, clay, gravel, culm, granulated slag, mill tailings, and even ashes, have been used as filling material in place of or in conjunction with rock. Of all the materials used for filling the best is probably clean sand. Gravel, finely crushed rock, granulated slag and coarse mill tailings are almost as good, and the methods used for placing and confining sand will apply equally well to any of these. The advantages of sand are its comparatively slight compressibility and the ease, economy and convenience with which it may be placed. On account of the voids, which necessarily occur in fills made of coarse waste, such fills will compress 10 to 25 per cent before they develop their full strength, whereas clean sand seldom compresses more than 5 per cent. Sand is especially desirable when practically complete filling is required, and sand derived from mill tailings, especially when not thoroughly cleaned, has been remarkably effective in sealing off and extinguishing mine fires. When natural sand is available on surface, or when it can be obtained from the mill tailings or tailings dump, the cost of using this material for fill will probably be considerably less than that of rock. In this paper it is not proposed to discuss all methods of filling, but to describe some of the methods used for placing sand fill and to make a brief comparison of their merits. Methods of Placing Sand Sand may be transported and placed underground by several methods, which may be classified as follows: 1. By Hand.—The sand, which must be comparatively dry, is taken down the shaft in cars on cages and is distributed to its destination in these cars, or it is dumped down raises and is drawn off and distributed by means of cars, wheelbarrows or scrapers.

Jan 1, 1932

By F. G. Coggin

" If you could only get that motion into a machine," said a gentleman, as he watched the process of making a " van " on a shovel, and saw the copper roll up to the highest point, "it would beat the world for slime-dressing." This idea has led to many an effort to get into a machine that peculiar motion, called the " vanning motion;" and so cams, eccentrics, eccentric gears, levers, bell-cranks, togglejoints, links, springs, etc., have been brought into requisition, while a contribution has been laid upon all the conceivable motions, rolling, tumbling, swinging, jerking, tossing, dropping, tripping, intermittent, regular and irregular, or a combination of some of them, to which have been added percussion, concussion, and more discussion —all in a vain attempt to combine in a single machine the consecutive motions of the shovel, in washing, flowing, and tossing the van. The Patent Office Reports, and the various mills of the country, show, as a result of this, a lot of gimeracks, most of which never went beyond a first trial, finding their way to the scrap-heap. Others still running might as well be there; while some have considerable merit for dressing other and richer sands than copper. It does look tempting to see that clean edge of copper climbing up the shovel-blade; but those who have seen in it such visions of large fortunes have not considered the time it takes to make a van, compared with the small quantity treated upon the shovel; and this points to the reason for the failure of all that class of machines called " vanning machines," when used for dressing copper slimes, namely, a lack of capacity. The writer has come to this conclusion through an experience with the Rittinger side percussion-table of the most improved make, sent from Germany to the Calumet and Hecla Mining Company, and with the Ellenbecker tables, used by the same company, both of which are good types of that class of machines. In both of these machines, the tables are suspendetl by rods which allow them to swing, and they can also be adjusted for the purpose of varying the inclination of the table. In the Rittinger machine, the motion of the table is at a right angle to the flow of the slime, being produced by a wiper-cam which throws the table against a spring, which throws it quickly back against a stop, the percussion throwing up the mineral. In the Ellenbecker table, the motion is

Jan 1, 1884

By Nigel B. Pearson

INTRODUCTION Chingola Division of Nchanga Consolidated Copper Mines Limited is situated in the Copper- belt of Zambia, approximately 30 kilometres south of the border with Zaire. It is the country's leading producer of copper and its principal foreign exchange earner, and has been, in its time, the largest single copper producer in the British Commonwealth and the second largest in the world. There are two 'primary' copper sources at the Division, namely Underground producing approximately 100 000 tonnes of contained copper annually, and the Open Pits, producing approximately 200 000 tonnes annually. Following the commissioning of the Tailings Leach Plant in 1974, treating final tailings, metallurgical recovery increased from approximately 70% to in excess of 80%. Extensive deposits of cobalt are present and following the rise in the market price in 1979 and 1980, plans were made to extract this separately. Cobalt production commenced in 1980 from Underground and will increase substantially when the Open Pit cobalt reserves become available in approximately 2 years time. This paper covers the development of the production method of the principal Underground Orebody, known as the Lower Orebody, and the control systems applied to it. HISTORY The Lower Orebody, formerly known as the Nchanga West Orebody, was discovered by extensive Diamond Drilling operations in the late 1920's. Development of the Shaft complex and the orebody commenced and continued until 1931 when flooding from development on the 480 foot horizon necessitated closure. The mine remained closed until 1937 and production commenced from the Lower Orebody in 1939. Various expansion schemes of the Metallurgical Plants and the Underground workings were completed over the years with the result that Underground production increased steadily to in excess of 150 000 tonnes copper annually in the 1960's. In recent years production has levelled out at slightly less than 100 000 tonnes annually. This decrease is as a result of producing from a lower grade reserve, available within a more limited production front, and the application of production constraints in the light of experience gained. GEOLOGY Within the Nchanga Mining Licence Area of Chingola Division, there are a total of 14 known copper orebodies in the basal part of a thick succession of sediments of Pre-Cambrian age. The sediments rest on a Basement Complex of granites, gneisses and schists. Three superimposed stratiform orebodies, known locally as the Lower, Intermediate and Upper Orebodies, occur over a vertical interval of approximately 150m. Current annual copper production is approximately 300 000 tonnes. The ore originates from four different mining operations, one Underground mine and three Open Pits. Approximately 9 million tonnes of ore grading between 3 and 3.5% copper are milled annually. The two largest sources of ore, the Under- ground mine and the Nchanga Open Pit are on the gently dipping south limb of the Nchanga Syncline (Figure 1). In the Nchanga Open Pit the major source of ore is the Upper Orebody with lesser tonnages from the Intermediate and Lower Orebodies. In the Underground mine the Lower Orebody is the only orebody being mined, underground mining of the Upper Orebody being suspended for economic reasons in 1978. The typical stratigraphic column is illust- rated in the legend on Figure 1. The footwall rocks comprise either Nchanga Red Granite or the Arkose formation, both of which are hard competent rocks, offering generally good mining conditions for footwall development. Copper mineralisation of the Lower Orebody extends from the upper part of the Arkose to the top of the Lower Banded Shales. Overlying the Arkose is the Transition zone (0-18111 in thickness). The Transition zone comprise a lower member, Transitional Arkose, consisting of incoherent kaolinised sandstone, and an upper member, Transitional Lower Banded Shales consisting of normally soft bleached and silicified shale. The soft nature of the Transitional rocks is

Jan 1, 1981

By Harvey T. Kennedy, Erasmo T. Guerrero

In this investigation, the effects of the surface tension of brine and the interfacial tension between oil and brine on the recovery of oil by water flooding of Woodbine sand were studied. Variation of the surface forces was obtained by the addition of small amounts of commercially available chemicals, which did not appreciably change the surface tension of the oil. Fourteen surface-active agents, which were found to be effective in lowering the surface and interfacial tensions, were further investigated to determine the extent of their adsorption by the Woodbine Sand cores. It was found that adsorption was sufficient in all cases to prevent any substantial concentration of chemical at the flood front. The flow system was, therefore, brought to equilibrium with definite concentrations in the flood water by circulation before each test. The data thus have no relation to the results obtainable by using the chemicals in full-scale water floods, but should be helpful in predicting the performance of surface-active chemicals which are not lost by adsorption, if such can be found. Flow tests were made on various concentrations 01 six chemicals selected from the 14 investigated for surface activity and adsorption, with and without the presence of gas in the core. It was found that lowering the interfacial tension had a tendency toward reducing the recovery of oil and that the reduction of the surface tension of the brine had little or no effect. A substantial beneficial effect, however, resulted from the presence of gas in the cores. INTRODUCTION The movement and distribution of crude oil in a porous medium are governed by such factors as surfacc and interfacial tensions, viscosity of the oil, pore size configuration, and wetting characteristics of the medium. The knowledge of the part played by each of these factors is limited. If the oil is assumed to be distributed in droplets whose movement is restricted only by the pore openings, the following equation may be applied: ?P= 2r(1/r1-1/r2) .......(1) In this equation ?P is the pressure drop necessary to force an oil droplet through the construction, r is the oil-brine interfacial tension, r1 is the radius of curvature at the forward and smaller end of the deformed droplet, and r2 is the larger radius. This equation neglects the effects of viscosity, as a static or near static condition is assumed. It would apply equally well for a bubble of gas in liquid if r were taken as the surface tension of the liquid. This relation shows that droplets can be more easily displaced if the interfacial tension is decreased. The displacement of oil from the surface of a sand particle may also be considered from the basis of selective wetting assuming that the oil droplets are smaller than the pore constructions and that the sand surfaces may be considered as planes. A relation of the forces involved can be derived from Dupre's equation to give: Wos = row (1 +Cos?)........(2) where Wos = Work of adhesion between the sand and the oil row = Oil-water interfacial tension 0 = Contact angle ros = Oil-solid interfacial tension This equation would indicate that the displacement of oil from a sand surface is more readily obtained when the water-oil interfacial tension is reduced provided the contact angle remains constant. Evidently. however, the values of Cos 0 may vary along with the water-oil interfacial tension due to changes in the other surfaces of the system. The above theoretical considerations suggest that perhaps even the static behavior of a dispersed oil phase in a porous medium is considerably more complex than may be inferred by the simple systems

Jan 1, 1955

By N. Brown, H. Green

The effect of quenching temperature and of aging temperature and time on compression stress-strain curves of ß brass was investigated. Age softening occurs at a rate which decreases with decrease of quenching temperature below the critical temperature for ordering. The activation energy for softening is 15,400 cal per mol. No matter how rapidly brass is quenched or what the temperature from which it is quenched, practically complete order is obtained at room temperature. It has been demonstrated that with respect to the specific heat,',' electrical resistivity," X-ray diffraction, and microstructure, the slowly cooled and rapidly quenched specimens are the same and that complete long range order exists in both cases. However, Smith" showed that the hardness of ß brass depended upon the cooling rate and the temperature from which brass is quenched. Smith showed that the hardness was a maximum for specimens quenched from about 445 °C and that the as-quenched structure softened at room temperature at a rate depending upon the temperature prior to quenching. This softening after quenching is in this paper referred to as age softening. The results of this investigation are qualitatively the same as the findings of Smith. However, a more detailed analysis of these phenomena has been possible since the stress-strain curve offers a more sensitive index of the state of the metal than hardness. It is the purpose of this paper to show how the mechanical properties are related to the ordering of ß brass. A 200-lb ingot made from electrolytic copper and electrolytic zinc was extruded to 1/2in. rod. Chemical analyses at frequent intervals along the rod showed that the alloy contained 51.5 pct Cu with the greatest impurity being iron, at 0.01 to 0.001 pct. To obtain a constant test material, the 1/2in. extruded rod was warm-rolled to % in. diameter at 200°C and then recrystallized to a grain size of 2.0 mm. The specimens for the compression test were % in. high and % in. in diameter. The specimen ends were polished on 600 metallographic paper and coated with Molykote for a lubricant. The compression rig consisted of an anvil resting on the machined table of the Baldwin Southwark testing machine and a plunger fixed to the movable cross

Jan 1, 1954

By Lyall Zickrick

Numerous brands of virgin pig lead are used in the manufacture of lead and lead-alloy cable sheath for commercial cables. In changing from one brand of pig lead to another a readjustment of press operating conditions is nearly always necessary. This is due largely to different flow characteristics, and some brands extrude at a much lower pressure than others; for example, copper-bearing lead requires much more pressure than a low-bismuth lead under the same operating conditions. With the modern type of lead press the extrusion practice in general differs very little. Molten lead at a temperature of 375" to 400" C. (710" to 750" F.) is cast in a water-cooled cylinder and allowed to cool or solidify into a slug for six to nine minutes, This slug is forced into the die block and around the core tube through which the cable runs. The • die block is held at an optimum temperature by means of gas flames, usually 175" to 225" C. (350" to 435" F.). Measurements of the temperature of the cable sheath taken at the time of extrusion have covered a range of 190" to 225" C. (375" to 435" F.). According to Tammann, the rate of extrusion for a metal under constant pressure through a simple orifice is almost doubled for each 10" C. rise in temperature. Because of the irregularity of die-block construction, such a great increase is not obtained in lead-cable extrusion practice. To obtain more definite information about the effect of small percentages of metals usually present as impurities in virgin pig lead, a series of compression tests were made covering several different brands of lead and various alloys. Since the usual temperature of the sheath is approximately 200" C. as it emerges from the die block, these compression tests were all performed at 200" C. Five different brands of virgin pig lead were selected, as follows: Brand A: A desilverized western lead of very high purity. This lead contains neit,her copper nor bismuth. Brand B: A bismuth-free southeastern Missouri lead, usually containing 0.05 to 0.08 per cent copper. Brand C: A copper-free mid-western desilverized lead of very good grade, usually containing around 0.06 to 0.10 per cent bismuth.

Jan 1, 1932

By R. I. Jaffee, C. M. Craighead, G. A. Lenning

The a-p type alloys are subject to a loss of tensile ductility with increasing hydrogen content. No hydride phase is visible in embrittled a-B type alloys. The embrittlement encountered appeared to be of the strain-aging type. Both compositional and structural factors are shown to influence the hydrogen tolerance of a-B type alloys. IN previous papers by the authors, the effects of hydrogen on the structure and mechanical properties of high-purity and commercial-purity titanium were described. It was shown that the presence of hydrogen in these materials results in a hydride phase having low solubility at room temperature. This insoluble phase had little effect on the room temperature tensile properties, but did decrease the notch-bar toughness to a large degree. The binary Ti-H systemz indicates that hydrogen is much more soluble in £ than in a-titanium. Hence, it is not expected that the embrittling effects of hydrogen in a-fi alloys will be similar to that found in a alloys. Materials of this type studied in the present investigation included two commercial a-/3 alloys, Ti-8Mn and Ti-4A1-4Mn; three high-purity Ti-Mn alloys containing 3, 6, and 9 pct Mn; and four high-purity Ti-Mo alloys, Ti-SMo, Ti-10.9M0, Ti-13.1 M0, and Ti-20M0, of the a-B and B types. Materials and Fabrication—The high-purity base alloys, composed of iodide titanium and high-purity metals, were prepared by double arc melting under argon in a water-cooled copper crucible, using a water-cooled tungsten electrode. Radiographic examination showed homogeneous Ti-Mo alloys. The analyses of the alloys used in this investigation and their as-fabricated hydrogen content, where determined, are listed in Table I. The two commercial alloys were 1/2 and % in. diam bar stock. No analyses other than hydrogen content were made. In order to maintain the a-B relationship existing in the two commercial alloys, vacuum annealing and hydrogenation were carried out in the B field prior to final fabrication. Fabrication then was done by The three high-purity Ti-Mn alloys were forged to 3/4 in. diam rods at 1600°F, and hot swaged to 1/2 in. diam at 1380°F. Machined and degreased bar stock was vacuum annealed or hydrogenated, fabricated to 1/8 in. diam rod by hot swaging, annealed in argon for 1 hr at 1380°F !75O°C), and water quenched. The six yz lb Ti-Mo ingots were forged to Wi in. diam rods at 1600°F and then hot swaged to % in. diam. Machined and degreased specimens were vacuum annealed or hydrogenated in the /3 field, prior to fabrication, by swaging to Y4 in. rod at 1400°F. After fabrication, the 5 and 10.9 pct Mo alloys were annealed in argon for 16 hr at 1290°F and water quenched to give equiaxed a-B structures. The higher molybdenum content alloys were argon annealed 4 hr at 1470°F and water quenched to give all £ structures.

Jan 1, 1957

By Chas. A. Dobbel

The Chief Consolidated properties are situated in the Tintic mining district of Utah, being included in Juab and Utah Counties, about 70 miles south of Salt Lake City. The drilling referred to in this paper has been carried on in the Chief Consolidated No. Pmine, the Gemini mine, and the Eureka Hill mine at Eureka, the Grand Central mine, above Mammoth, and the Apex Standard mine just south of Dividend. The occurrence of ore through the district is typically in lenses on bedding or breaks, and in connecting pipes, without accurate geometrical trend so that the ratio between the total length of development in drifts and crosscuts to the square feet of ore exposed is very high. A section of one of the western channels (Fig. 1) through the Tintic district, shows, roughly, the irregularities calling for excessive development work, so typical of the Tintic mines. This irregular trend of the ore constantly calls for excessive cutting up of the whole area for offshoots and the connecting links always present. In the planning of prospecting and development, no regular system may be laid out without overlooking some orebodies or segments of orebodies. It was with these facts in mind that Cecil Fitch, the general manager of the Chief Consolidated Mining Co., decided that some method of drilling was the more feasible way of successfully cutting up and bringing about the discovery of ore believed to exist; yet this drilling had to compete with a great margin of cost over crosscutting, as a crosscut in the Tintic district, even though barren, brings out geologic facts that, in many cases, are worth the price of the working itself. Diamond drills met with the success that might be expected in a district that has a reputation for the Gear and loss of diamonds; after 6 or 8 months, this method was discarded on account of the high costs. Simultaneously, a small churn drill outfit was set up underground but a few shifts proved its inapplicability. At the same time, the ordinary hammer drill was used to drive prospect holes with jointed steel. Just what equipment to develop for this work with the hammer drill involved much experimenting, as nothing was found in the literature

Jan 1, 1925

By Nelson Severinghaus

Crushed stone leads all industrial mineral, except cement in annual value of production in the United States. It is also interesting to note that crushed stone value is exceeded only by that of fuels, cement, and iron ore among all mineral products of this country. An industry which started with hand-sledging to produce smaller sizes of broken stone for hard surfaced improvement of roadways as designed by McAdam in England about 1815 has grown into a complex industrial giant. Most of this growth started with the invention of a mechanical rock crusher just under one hundred years age (1858) by Eli Whitney Blake. This invention made possible the production of greater tonnages of smaller sizes of broken stone and greatly accelerated its use for Macadam road pavements and for railroad ballast. Stones Used for Crushing All three of the major geologic classifications of stone, sedimentary, igneous, and metamorphic, are widely used in the production of crushed stone. Satisfactory physical and chemical properties can be found in each type so that economic availability becomes the prime determinant for development. Table 1 lists more common rocks of the three main groups with average values of specific gravity and toughness as measured by Deval test. This widely used classification was suggested by C. E. Lord, of U. S. Bureau of Public Roads. The physical properties were compiled by D. O. Woolf in "The Results of Physical Tests of Road Building Rock," U. S. Department of Agriculture, Bulletin 76. It should be noted that the figures given in Table 1 are averages. Very substantial variations in the physical properties of rocks within any one classification will occur. The solid surface of the earth was first made up of igneous rocks resulting from the solidification or crystallization of molten material upon cooling. These igneous rocks were disintegrated into sand, clay, and gravel by the weathering action of wind, water, chemicals, [ ] and temperature changes. The products of weathering were transported by wind and streams until they reached less turbulent areas such as oceans or lakes and were deposited. Chemical deposition between grains of sand or clay or the physical action of pressure and heat then solidified them into sedimentary rocks

Jan 1, 1960