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By M. Gensamer, B. Lustman

A recent paper described, by means of stereographic pole figures, the preferred orientations produced by cold-rolling low-carbon steel of automobile-body grade. It was found that the surface of the sheet exhibited a deformation texture significantly different from the texture possessed by the inside of the sheet. Efforts to produce the surface texture by buffing a sample exhibiting the normal or inside texture, using a copper lap, have not been successful. In view of the data collected by Kurdjumow and Sachs2, it was suggested that the surface texture might be the result of partial recrystallization of the normal or inside texture. The recrystallization texture of cold-rolled mild steel was studied by Goss3, but he did not use the pole-figure method, so that his conclusions were in need of verification. For these reasons the work reported in here was undertaken. It had two objectives: (1) to study the general behavior of cold-rolled mild steel on recrystallization below the critical temperature Al, and (2) to attempt experimentally a demonstaration of the genesis suggested above for the surface texture in cold-rolled steel. The material was that used by Gensamer and Mehl, cold-rolled to 97.5 per cent reduction in thickness. Fig. 1 is an X-ray pinhole photo-gram of this material, taken with the beam transmitted perpendicular to the plane of the sheet, using a molybdenum target at 35 kv. with no filter. It is typical of the inside texture, but in it the deformation texture is better developed than in the samples that showed a well developed difference between the inside and surface samples. Specimens from this sheet, which was 0.005 in. thick, were annealed for 5 min. at various temperatures in a lead bath. After about 0.0005 in. had been etched from both surfaces, X-ray photograms were prepared, using the technique described in the earlier paper1. Figs. 2 to 6 are photograms of the annealed sheet, at 566, 571, 577, 580 and 650" C. (1050' to 1202" F.), in order. Fig. 2 shows that 5 min.

Jan 1, 1937

Lucien Eaton, Ishpeming, Mich. (communication to the Secretary *):—I began to use water Leyner drills at the Cliffs Shaft mine at Ishpeming, Mich., over four years ago, and now have over 20 in use. The ore in this mine is hard specular hematite, and is mined by the room-and-pillar system, the rooms being driven as breast stopes. For many years only 3 1/4-in. piston drills were used in the mine, each machine ordinarily being able to drill from 15 to 25 ft. of holes in a shift. In the hardest ground, however, it was often, impossible to drill more than 2 or 3 ft. in a shift. When water Leyner drills were used the footage drilled was materially increased, especially in the hardest ground, but it was found that in very hard ground the breakage of machine parts, and especially of drill steel, was very high. At this time round steel with lugs forged on the shanks was used, and most of the breakage occurred at or near the lugs. Many experiments were made in an endeavor to overcome this breakage, and it was finally decided to do away with the lugs altogether, and use hexagonal steel. Accordingly I designed a chuck to use hexagonal steel without luge, a tappet being interposed between the hammer and the shank of the' drill steel. The design of this chuck of the proper size for the No. 18 Ingersoll-Leyner drill is shown in Fig. I. The different parts are also shown photographed in Fig. 2. A is the barrel of the chuck, which revolves inside the front head. It is made of machinery steel. B is the bushing which screws into the front end of the barrel A, and in which the drill steel fits loosely. This is turned down from a steel casting, the hexagonal hole being cast and then broached out to the proper size. It is case hardened inside. C is the tappet, which fits smoothly inside A just behind the bushing B. It is made of " N " temper Crescent tool steel and is carefully tempered. D is the tappet ring, which prevents the tappet from being pushed back too far by the

Jan 1, 1914

By R. W. Gurry

IN the course of a series of measurements of the rate of diffusion of carbon in austenite at about 960°C. (1760°F.) and 1110°C. (2030°F.), it became necessary to determine carbon concentration when austenite is saturated with graphite, because of indications that the hitherto accepted value of the solubility of graphite in this temperature range is too low. Specimens of carbonyl iron were saturated at temperature in a hydrogen-toluene atmosphere which was in equilibrium with graphite. The carbon content of the steel was measured either by combustion or by the loss of weight after substantially complete decarburization in an atmosphere of hydrogen saturated at room temperature with water vapor. The absence of carbide or graphite as a phase at temperature was demonstrated by microscopic examination of saturated specimens that had been drastically quenched. The mean result of concordant measurements is that gamma iron saturated with respect to graphite contains 1.39 per cent carbon at 957°C. (1755°F.) and 1.89 per cent at 1110°C. (2030°°F.). These two values, plotted along with what appears from the literature to be the best value (0.69 per cent) at the iron-graphite eutectoid temperature (738°C. or 1360°F.), yield a curve that is almost linear, and which, continued upward through a short interval, leads to a limiting solubility of 1.98 per cent at the iron-graphite eutectic temperature1135°C. (2095°F.) instead of the commonly accepted value of about 1.7 per cent. EXPERIMENTAL PROCEDURE The samples of carbonyl iron, with the exception of those for microscopic examination, were cylinders 3.8 cm. long, 0.475 cm. in diameter. They were placed in a graphite holder and suspended in a vertical furnace, the temperature of which was controlled precisely by means of the resistance of the platinum winding.1 In the empty furnace the range of temperature over a distance of 4 cm, at the heat center was 9°C. (16°F.) but at a given location the temperature remained within ±0.5°C. (0.9°F.); it was measured by inserting a platinum-rhodium couple calibrated at the melting point of palladium, 1555°C. Since the presence of the sample probably improved the uniformity of temperature, it was decided to consider the average temperature of the sample to be 3°C. lower than the maximum measured in the empty furnace with the same setting of the controller. Data were obtained in the vicinity of two temperatures, 960°C. (1760°F.) and 1110°C. (2030°F.). The samples for microscopic examination, 2.2 cm. long, 0.475 cm. in diameter, were suspended individually on pure iron wire; for these no temperature correction was applied. The carburizing gas passed into the furnace was dry, oxygen-free hydrogen saturated at room temperature with

Jan 1, 1942

By Ellsworth Daggett

Jan 1, 1879

By Brandon H. Grove

In Austria's only producing field, at Zistersdorf, Erdoel Produktions-gesellschaft completed a single new producer, its well Gosting 6 coming in early in October for an average production of about 15 tons daily. For further testing within the proved part of the field this company's Gosting 7 was at the close of the year drilling in Pannonian strata at 760 m., and its Gosting 8 was rigging up. Outside of the proved area of the Zistersdorf field two further tests located on the Steinberg dome failed to find production in the known oil-bearing strata of the Sarmatian, and were deepened as wildcats into the Cretaceous Flysch. The first of these, Neusiedel 1, about one kilometer north of the known producing area, found oil in the Flysch beds at a depth of 1290 m. for an initial production of about 10 tons per day. The second test, Windisch Baumgarten la, is an old well previously considered a failure, which when deepened achieved a production of about 5 tons per day in the Flysch Inoceramus beds at 985 m. Neither of the foregoing wells is producing from a stratigraphically definable horizon, oil accumulation in the Flysch of this area being apparently controlled by tectonic breccias or related crevices.

Jan 1, 1937

By James Norman

SPODUMENE is a lithium-bearing pyroxene, and is an important source of lithium compounds. Because of its high alumina and lithia content it might be a desirable constituent of glass batches. The use of spodu-mene in ceramics will undoubtedly increase, as it has been shown that a mixture of the proper proportions of this mineral with soda feldspar and potash feldspar has a much lower maturing temperature upon firing than has any one of the three components.1 The mineral spodumene seldom occurs as pure mineral, but ordinarily with feldspars, mica and quartz. All spodumene-bearing rocks tested contained these minerals as gangue. Small amounts of iron and manga-nese oxides also were present, usually as stains on the silicate grains. Mica associated with the spodumene carried a considerable amount of lithia. This was discovered in flotation tailings, which, although virtually free of spodumene grains when examined microscopically, showed 0.4 to 0.5 per cent Li20 on chemical analysis. The lithia content of micas picked from the North Carolina spodumene rock was variable. Some samples contained little lithia, others more than 3 per cent. In 1938, under the supervision of Oliver C. Ralston, ? the Bureau of Mines, in cooperation with the Tennessee Mineral Products Corporation, undertook to study the beneficiation of North Carolina spodumene by froth flotation. The Bureau had previously published investigations on the beneficiation of spodumene.2,3 The methods described in the second publication' did not give good results with badly weathered spodumene rock, so that more thorough investigation of the subject was warranted.

Jan 1, 1940

It is not often that a mining district is discovered, opened up, exploited and developed into a rich producer within the space of three or four years. Not until the year 1903 was the existence of the rich cobalt-nickel arsenides and silver of Northern Ontario even suspected by the mining public. But it is worthy of record that Dr. W. G. Miller, the man who has done so much to develop sanely and honestly the mineral resources of Ontario, pointed out [ ] the possibilities of this region at least two years before the first discovery was made. . Dr. Miller, then Professor of Geology at the School of Mining, Kingston, Ont., referred in these words to the district: "It will be seen from what has been stated on preceding pages that the district examined contains as great a variety of rocks as probably any part of the Province of equal area . . . Although few discoveries of economic minerals have been made in this territory, it may reasonably be expected that deposits of value will be found when

Jan 1, 1907

By J. O. Betterton, W. M. Spicer

The solubilities of antimony in the a and ß-forms of zirconium were determined between 750" and 1300°C. The a-phase forms by a peritectoid reaction, and the a/(a + ß) and ß/(a + ß) phase boundaries rise in temperature with in- ' creasing antimony. Impurities in the 99.95 pct zirconium were sufficient to influence the phase boundaries and the variation reported is approximate. The restricted solid solubility of the terminal zirconium phases, which is especially characteristic of this system, is thought to result from the low free energy of the Zr2Sb phase. ZIRCONIUM-antimony alloys are important in zirconium alloying studies since antimony is one of the highest valent elements which can be used to make zirconium alloys. The dilute alloys are of further interest as they have somewhat better corrosion resistance' than pure zirconium. The first investigation of the phase diagram was made by Russi and wilhelm2 who found a eutectic reaction, liquid (- 17.5 at. pct Sb) = ß zirconium (- 11 at. pct Sb) + Zr2Sb, at 1430°C'. The Zr2Sb phase was of complicated hexagonal structure, and by thermal analyses of the relatively impure alloys then available, Russi and Wilhelm found evidence for the peritectoid formation of the a-phase. EXPERIMENTAL In the present work, the phase diagram was re-investigated in the region 0 to 5 at. pct Sb and 750" to 1300°C using iodide-zirconium alloys. Details of the procedure have been described earlier. Essentially the alloys were prepared by arc casting and were then examined by the annealing and quenching method. The times of annealing varied from 1 hr at 1300°C to 30 days at 760°C. As in other zirconium phase-diagram studies1"5 the impurities in the zirconium were still sufficient to affect the results, and the experiments will therefore be described in 3 parts. The results from the first 2 parts of the work are shown in Fig. 1. Here the ß/(ß + Zr2Sb) and a/(a + ZrzSb) phase boundaries were determined with alloys made from 99.9 pct zirconium,* and cast under such conditions that the Vickers hardness of a alloys extrapolated to 95 to 110 VPN at 0 at. pct Sb. Such alloys were not suitable for investigations of the (a + ß) region, and the results shown in Fig. 1 near the (a+ ß) and (a +ß + Zr2Sb) regions were determined with purer alloys. These alloys had an extrapolated hardness range of 75 to 85 VPN and were made from 99.95 pct zirconium. A deviation from binary equilibrium occurs in Fig. 1 as indicated by the (a + ß) region of pure

Jan 1, 1959

By Max Kraut

UNTIL VERY RECENTLY the flotation process has not found much application in the treatment of gold ores. No appreciable improvement has been made lately in the technology of this application; but the possibilities of flotation have been more fully appreciated. Excellent results were obtained long ago, but few operators paid much attention to them. For instance, in 1915 extensive experiments were carried out by Mr. Beauchamp and Minerals Separation on ores from the Dutch-Sweeney mine near Jimtown, Calif. The ore contained gold-bearing pyrite and laboratory tests indicated a 10c. tailing on a $6 head. In 1918 a 300-ton flotation mill using Minerals Separation machines was built based on the results obtained. However, when the mill began full-scale operation, mill heads did not come up to expectation, being only $4; also the-tonnage was considerably less than had been expected. Flotation mill tailing assayed 15c., indicating a recovery of over 96 per cent, so that excellent metallurgical results were obtained. The operation was a success but the patient died.

Jan 1, 1932

By A. K. Chakravarti, A. Lihiri, G. G. Sarkar

One primary objective of India's third Five Year Plan is the expansion of coal production. The goal: boost present output of 45 million tons to 95 million tons by 1965--an increase of more than 100 pct!

Jan 7, 1961

By D. E. Norquist, J. E. Moeller

The results of wear on marked balls, 4, 31/2, 3, and 2 in. diam are given. All balls were forged steel of practically the same chemical analysis and hardness. The results indicate that balls in a given mill for a given length of time will have equal diameter losses, regardless of size.

Jan 6, 1950

By C. Richard Tinsley

INTRODUCTION Project financing has been defined by various authors in this book. It is founded on reliance upon the project's own future cash flows and secondarily on the collateral value of the asset base developed. It is chiefly provided by the banking community who, after an assessment of the risks involved (See Chapter 9) analyse the company's presentation (See Chapter 10) in order to structure a financing package often at a leverage of up to 80% debt. This paper describes the various support and structuring techniques used to delimit and abate the risk components perceived from the banks' experience on project development especially for the period before cash flow is generated (pre completion). ADVANTAGES The advantages of project financing extend beyond the offloading of risks to banks. Among the advantages are the following: 1. Longer loan maturity than conventional, unsecured financing; 2. Loan disbursement and repayment tailored to specific mine; 3. Ability of a joint venture without a "track record" to raise financing; 4. Obtain financing where further direct borrowing (corporate credit) limited by existing indentures; 5. Possibility of arranging finance off-balance sheets; 6. Leverage return on equity; 7. Obtain tax savings with certain project financing structures; and 8. Obtain flexible loan repayment conditions. It can also be said that from the point of view of the company, project financing has certain disadvantages which can be summarised as follows: 1. Perceived higher cost (risk premium) compared with straight corporate credit; 2. More documentation; and 3. Need to negotiate risk-sharing aspects. The main attraction of project financing is the ability to pass certain risks on to the lenders within a flexibly funded transaction. (For a more detailed description of the specific risks see the author's paper at the beginning of Chapter 9. The same risk system is followed here.) The typical project financing has corporate support prior to completion and will usually carefully control the resultant cashflow after completion to ensure the bank does not unnecessarily assume further risks of an equity nature. WITHIN THE COMPANY'S CONTROL Operating Risk This is broken into three interrelated segments for ease of description. Technical Component: This must be specifically addressed if new technology or unproven adaptations are to be used. The banks will have scrutinized the sampling and bulk testing work, perhaps grilling the geological, mineral processing, metallurgical, and mining engineering staff or consultants directly. A number of supports are available such as: TECHNOLOGY GUARANTEE: Either the company itself or the process developer will guarantee that the technology will work for a period of, say, six months to two years after startup. If not, then financial penalties will apply although these are rarely sufficient to repay the loan. Many process developers are weak financially and may have little substance to back such a guarantee.

Jan 1, 1985

By Harold Vagtborg

Equilibrium vaporization ratios were obtained for light hydrocarbons, nitrogen, carbon dioxide, and hydrogen sulfide in a reservoir fluid containing 35 mole per cent hydrogen sulfide. The data cover the range from 700 to 2500 psia at 154°F. The results of this study are compared with results obtained by other investigators on systems containing less hydrogen sulfide. When large amounts of hydrogen sulfide are present, the equilibrium vaporization ratios for ethane and heavier hydrocarbons are greater than the ratios for systems containing little or no hydrogen sulfide. Large amounts of hydrogen sulfide have the reverse effect on the ratios for methane. Routine PVT studies were performed on the reservoir fluid including flash and differential liberations of the dissolved gas. The pressure-volume relationships of the saturated fluid were also determined. It is shown that the use of published correlations on physical properties of reservoir .fluids can lead to gross errors when applied to systems containing large amounts of hydrogen sulfide. INTRODUCTION Equilibrium vaporization ratios for various components in naturally occurring crude oil and condensate systems have been reported in the literature.1,2,3,4 Recently, several papers have been published with emphasis on non-hydrocarbon components such as nitro- gen, carbon dioxide and hydrogen sulfide.5,6,7,8,9 The increased discovery of reservoir fluids having high concentrations of these non-hydrocarbon components, particularly hydrogen sulfide, has created a need for a further study of their phase behavior. Equilibrium vaporization ratios are known to vary with pressure, temperature, and overall composition of the system.'" It was anticipated that ratios for systems containing large amounts of hydrogen sulfide would differ from the ratios for systems containing small amounts of hydrogen sulfide. Therefore, the present work was undertaken in order to obtain reliable equilibrium vaporization ratios for the reservoir-fluid in question. PROCEDURE The procedure followed was evolved after considerable experimentation on methods of handling the very sour crude. It was found that drying agents (e.g. CaCl2 and CaSO4) absorbed an appreciable amount of hydrogen sulfide from the gas used to prepare the recombined samples. The concentration of hydrogen sulfide dropped as much as 10 mole per cent in a single pass through a 3-ft drying tube. It was also found that when undried gas was agitated with mercury in the PVT cell, a black scum, believed to be mercuric sulfide, was formed and the hydrogen sulfide content of the gas dropped several mole per cent. Because composition changes during a run could not be tolerated, separation of the mercury and sour fluid was necessary. In order to effect this separation, a stainless steel piston, equipped with Teflon O-rings, was placed in the PVT cell between the mercury and the sour fluid. This piston had to be constructed in three parts because

Jan 1, 1955

By AIME AIME

FOUK fields of activity are now recognized by the A.I.M.E. in its award of medals for conspicuous achievement: the Saunders medal for mining, the Douglas medal for non- ferrous metallurgy the Lucai medal for petroleum, and the Rand medal for distinguished achievement in the broad field of mining administration including the organization and development of enterprises in all branches of the mineral industry. This last is known officially as the Charles F. Rand Memorial Medal, and it award is made possible through funds supplied by the Charles F. Rand Foundation. Until this year money thus available was used for other special projects, so that the forthcoming presentation of the Rand medal to Robert C. Stanley is notable as being the first time that achievement in mining administration has been thus recognized.

Jan 1, 1940

By R. H. Wightman

The following paper describes current mining practices of the Riverside Cement Co. at its Crestmore plant, Riverside, California. For a number of years the Riverside Cement Co. obtained its raw materials by surface quarry operations but after these deposits were exhausted for equipment available at that time, the company started underground mining. Mine workings were begun in 1927 to 1930 for the recovery of limestone by block-caving methods. Geology The bed of limestone that is being worked is a metamorphosed and recrystallized limestone, which is generally white in color and from medium to coarsely crystalline. This bed is fissured by water courses that vary from minute cracks to large areas. The bed has also been cracked by intrusions along the footwall and hanging wall. These walls are coarse-grained, gray, quartz granodiorite. The strike of the deposit is north and south and it dips to the east at an angle of approximately 50°. The bed is approximately 1700 ft. long, 270 ft, thick at right angles to the dip at the middle of the deposit and tapers to a thickness of approximately 100 ft. at each end. Mining Methods The mine layout consists of four main levels with elevations as follows: 892 Collar of shaft. 800 Manway level used for ventilation, prospect purposes and as an entrance into the stopes. 700 Mining level where all material is extracted and on which are located the tapping grizzlies. This is the elevation from which all main cutoff stopes and undercutting work is started. 660 Haulage level where the limestone is drawn from transfer raises into the cars and is transported to the tipple by electric locomotives. 633 Level where all water is collected and which keeps the general water table below the working levels. A detailed description of the mine was given in U. S. Bureau of Mines Information Circular No. 6795, in June 1934. Current Block Development The first blocks to be caved were developed very much after the procedure followed in the copper mines of Arizona. This paper covers the ninth block to be caved. with a hard, tough rock like our limestone, it was found that some alterations in the development and control of dm, i, caving were. necessary to secure better crushing effect and provide more safety for such secondary blasting as is required. Sequence of Block Development 1. The fringes of the 800-ft. manway level are driven to confirm the diamond-drill information and serve to determine the final location and development required for the block (Fig. I).

Jan 1, 1946

By R. H. Wightman

THE following paper describes current mining practices of the Riverside Cement Co. at its Crestmore plant, Riverside, California. For a number of years the Riverside Cement Co. obtained its raw materials by surface quarry operations but after these deposits were exhausted for equipment available at that time, the company started underground mining. Mine workings were begun in 1927 to 1930 for the recovery of limestone by block-caving methods. GEOLOGY The bed of limestone that is being worked is a metamorphosed and recrystallized limestone, which is generally white in color and from medium to coarsely crystalline. This bed is fissured by water courses that vary from minute cracks to large areas. The bed has also been cracked by intrusions along the footwall and hanging wall. These walls are coarse-grained, gray, quartz granodiorite. The strike of the deposit is north and south and it dips to the east at an angle of approximately 50º. The bed is approximately 1700 ft. long, 270 ft. thick at right angles to the dip at the middle of the deposit and tapers to a thickness of approximately 100 ft. at each end. MINING METHODS The mine layout consists of four main levels with elevations as follows: 892 Collar of shaft. 800 Manway level used for ventilation, prospect purposes and as an entrance into the stopes. 700 Mining level where all material is extracted and on which are located the tapping grizzlies. This is the elevation from which all main cutoff stopes and undercutting work is started. 660 Haulage level where the limestone is drawn from transfer raises into the cars and is transported to the tipple by electric locomotives. 633 Level where all water is collected and which keeps the general water table below the working levels. A detailed description of the mine was given in U. S. Bureau of Mines Information Circular No. 6795, in June 1934. Current Block Development The first blocks to be caved were developed very much after the procedure followed in the copper mines of Arizona. This paper covers the ninth block to be caved. With a hard, tough rock like our limestone, it was found that some alterations in the development and control of draw in caving were necessary to secure better crushing effect and provide more safety for such secondary blasting as is required. Sequence of Block Development I. The fringes of the 800-ft. manway level are driven to confirm the diamond drill information and serve to determine the final location and development required for the block (Fig. I).

Jan 1, 1944

By L. G. Jones

An approximate method of calculation is developed in this paper which allows duplication of radial unsteady-state gas flow computer results where Darcy's law applies, such as those reported by Aronofsky and Jenkins and Bruce, Peaceman, Rachford, and Rice. Moreover, the new calculation method can be used to obtain results for radial unsteady-state gas flow obeying the quadratic flow law proposed by Duwez and Green. Means are discussed for predicting well behavior at single or superimposed flow rates in finite or infinite reservoirs, determining reservoir rock properties from well-test data, reproducing and interpreting back-pressure test data, and determining the radial extent and reserves of gas reservoirs from well-test data. Example calculations are presented for gas flow following both Darcy's law and the quadratic flow law. INTRODUCTION Since the publishing of U. S. Bureau of Mines Monograph 7,1 most gas-well testing methods have been based on the equation q = y (pl2 - pw2)a, where q = production rate, pl and pw, are formation pressure and sandface pressure, respectively, and y and a are constants to be obtained from test data. These methods, used for predicting both deliverability and "open-flow" capacity of gas wells, have been useful and accurate in many cases but unsatisfactory in others. Even at best, however, they do not supply information about the formation or lead to an understanding of nonsteady-state gas flow in porous media. Many theoretically based studies2-4 of gas flow obeying Darcy's law have been made.' Since the partial differential equations which result from combining Darcy's flow law with the continuity equation are nonlinear, most of the published research consists of either numerical solutions or analytical solutions for linear approximating equations. Such solutions have been of limited value in field work due to their unhandy form and their failure to correlate most field data. There is evidence5,6 which indicates that Darcy's law is inadequate to describe gas flow at some flow rates of practical interest. A quadratic flow law: which reduces to Darcy's law at low rates, is more successful in accounting for experimentally observed behavior. This flow law has been applied successfully to a few hypothetical reservoir cases in work which has recently been published.10 However, these numerical solutions of the equations involved have been successful only on a special analog computer. Routine applications to field cases would be awkward and have not been attempted. The present paper describes an approximate method for computing nonsteady-state gas flow solutions which has been completely successful in predicting the results for both Darcy flow and quadratic flow obtained by elaborate numerical methods. The new calculation method allows determination of the observable variables in gas-well testing at constant rates. It is similar to the scheme of using a succession of steady states suggested by Muskat7 in that it makes use of steady-state and material-balance equations. It also is similar to the work of Aronofsky and Jenkins2 in that the new method includes Aronofsky and Jenkins' stabilized flow equation as a special case. It improves upon both of these calculation schemes in that it accurately describes all portions of reservoir history and suggests means of determining reservoir rock properties from well-test data. This paper deals only with production from a reservoir, in which case the rate is defined as being negative. The reservoir studied here is a homogeneous, disk-shaped porous body of uniform thickness, with all boundaries sealed except the inner radial boundary of the well.

J: H. POLHEMUS, New York, N. Y. (written discussion*).-The Joplin .mining and milling practice has largely influenced operating methods in the Wisconsin district. Milling equipment is essentially of the same type with the notable exception that no tables are used. The treatment of the flue 'sands has been attempted several times but without commercial success. Table material coming from any one mill is low both in tonnage and zinc content. The iron is comparatively high as the pyrite-line "chat" is much more prevalent than the zinc-lime attached particle. In the finer sizes the zinc occurs mainly as free mineral. These condi-tions make the production of good-grade table concentrates very diffi-cult. The accompanying tabulated results from several mill tests indi-cate the character of the mill tails. I think the average mill tails in the district assay nearer 1.5 per cent. zinc than under 1 per cent., as stated by Messrs. Boericke and Garnett.

Jan 10, 1919

By Chester O. Ensign

Results of the application of geologic knowledge to grade control have been remarkable. Grade reduction from dilution has been diminished from greater than 9% at the time of the program's beginning to about 2% at present. This has resulted in an increase of about 1.8 lb of copper per ton of ore mined and has made a substantial contribution to earnings at White Pine during the past two years. Even more improvement is expected in the future. The White Pine orebody, situated in the southwestern portion of the Keweenaw peninsula of upper Michigan, is a gently dipping stratiform or bedded deposit. Copper distribution is stratigraphically controlled with higher values being confined to specific beds in the ore column. Room-and-pillar mining is practiced. Mining technique and geologic characteristics present a special array of problems in controlling the grade of ore.

Jan 4, 1964

By C. Schlumberger, M. Schlumberger

In order to explore electrically a terrain composed of a succession of horizontal beds, a current of known intensity i is caused to flow between two grounds A and B, and the resultant drop of potential , caused by the resistance of the earth, is measured between two given points in the soil, M and N. In practice it is convenient to choose the two points M and N so that they lie on the line AB and their midpoint coincides with the middle 0, of that line, although this is not imperative. When the lengths L and I of the two lines AB and MN are varied, equations are obtained which may be solved for the resistivity p of the subsoil at various depths. Recall briefly that the formula 4 x l gives the average resistivity, in the horizontal direction, of a layer of earth which extends to a depth1 about equal to L/4. This elementary concept is adequate for numerous practical applications. Nevertheless, it must be said that, in many cases, a rigorous mathematical analysis may be made, provided it is always a question of homogeneous, superposed horizontal beds. The authors expect to publish, later on, appropriate methods of calculation for some specific cases.

Jan 1, 1932