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RICHARD R. HICE, Beaver, Pa. (written discussion *).-The matter of selective. mining is probably of more importance in Pennsylvania than is washing, and perhaps washing would not be necessary at some plants where now practised if proper care were taken in the mining. I have in mind one plant where the sulfur content of the entire bed is more than 2 per cent., yet if the bottom 12 in. and the top 12 or 15 in. are separated, the main portion of the bed, some 6 ft., will not contain more than 1.25 per cent. sulfur; and yet in this mine the entire bed .is worked together, the resultant coke, as a matter of course, not being fit for use in iron smelting. Not sufficient notice is taken of the "Double-thick" Upper Freeport coal in Allegheny and Butler counties. This is a very valuable deposit. Other areas where there is a, top member of the Upper Freeport coal are also known, which add to the value of this bed as a coking coal. This double Freeport coal is especially suited for byproduct coking. I cannot be as optimistic as regards the "low-sulfur" Pittsburgh coal in Greene county as the authors. If the term low-sulfur is to he confined to coal with not exceeding 1.25 per cent. sulfur, there is probably little

Jan 10, 1919

By Kenneth Cottingham

THE area described herein is situated on what is known as the Arches Fork anticline in Roane and Calhoun Counties, W. Va. When the area was first mapped, it was felt that well records would give a more intimate idea of conditions, both structural and lithologic, than could be expected from surface horizons. The surface formations are chiefly members of the Monongahela series, though northward the hilltop outcrops are those of the Dunkard. Neither series offers a key horizon sufficiently definite to be satisfactory, although the structure as determined from well records was later checked against two of the more persistent surface sandstones. In mapping the subsurface structure, the top of the Big Lime (Greenbrier) was employed as the key horizon. The average depth to the top of the Big Lime varies from 1600 to 2000 ft. (487 to 609 m.), while the average thickness as indicated by well records is 113 ft. (34 m.). The oil-producing formation is the Big Injun, lying just below the Big Lime. In two instances, well records show the Lime and Injun separated by a few feet of "break." The Big Injun is recorded as having a variable thickness. Seven logs report it as missing and one gives it the maximum thickness of 105 ft. (32 m.), while the entire average for thickness is 19 ft. (6 m.).

Jan 1, 1922

AARONS, J. BOYD, Major. AGNEW, FRANK VANS. ALLEN, HERMAN H., Private, D Co., 8th -Reserve Engineers. BAINS, THOMAS M., Major, Engineer Officers' Reserve Corps. BANKS, H. R., Lieutenant, Canadian Engineers, B. E. F. BARBOUR, PERCY E., Deputy Supt. (Captain), New York State Troopers; Captain, 22d Regiment, N. Y. N. G. BARLING, H. B., Ist Lieutenant', Engineer Officers' Reserve Corps. BIGELOW, BRAXTON. BLOOMFIELD, E. C., Lieutenant. , BORROW, F. K. BROCK, R. W., Major, 19th Reserve Battalion of Canadians. BROOKS, ALFRED H., Captain, Engineer Officers' Reserve Corps. BROWNE, DE COURCY, 1st Lieutenant, Ordnance Dept., Fortress Monroe, Va.

Jan 9, 1917

(Members are urged to send in for this column any notes of interest concerning themselves or their fellow-members.) Members and guests who registered at Institute headquarters during December: William M. Peschel, Bradford, Mass. Samuel T. Wellman, Cleveland, Ohio. Herbert B. Cox, Boston, Mass. Samuel S. Wyer, Columbus, Ohio. C. Q. Schlereth, Mexico, D. F. H. H. Stoek, Urbana, Ill. H. W. Caldwell, Oxford, N. C. E. Harms, Torreon, Mexico. Charles E. Schwarz, St. Louis, Mo. G. A. Roush, So. Bethlehem, Pa. Dwight Furniss, Guanajuato, Mex. M. K. Rodgers, Los Angeles, Cal. F. W. DeWolf, Urbana, Ill.

Jan 1, 1914

By J. L. Taylor, Pol Duwez

THE present knowledge of the Ti-Al system is limited to the portion of the diagram extending from pure aluminum to the intermetallic compound TiAl3' A preliminary investigation of the titanium-rich Ti-A1 alloys revealed that the solubility of aluminum in titanium is quite extensive and that the transformation temperature of titanium is raised by alloying with aluminum. From an X-ray diffraction study of alloys containing up to 75 atomic pct Al (TiAl3), the phase boundaries at 750°C were located as follows: the a titanium solid solution extends from 0 to about 36 atomic pct Al, a two-phase region exists from 36 to 46 atomic pct Al, and an intermediate phase of varying concentration extends from 46 to 62 atomic pct Al. The purpose of the present paper is to describe the crystal structure of this new phase. The alloys were prepared by melting in a helium arc furnace on a water-cooled copper plate, using a furnace essentially the same as that described in ref. 2. The titanium metal was of the iodide type furnished by the New Jersey Zinc Co., and a piece of pure aluminum (99.99 pct) was obtained through the courtesy of the Aluminum Co. of America. After melting, the samples, weighing approximately 5 g, were sealed in evacuated fused silica tubes and homogenized for 4 hr at 1000°C. The specimens were then held for 10 days at 750°C and rapidly cooled to room temperature. Filings were then taken from each sample, sealed in evacuated vials, annealed at 750°C for 4 hr, and rapidly cooled to room temperature by quenching the vials in water. Powder diffraction patterns were obtained with a 14.32 cm diam camera, using K copper radiation filtered through a nickel foil. Structure Determination The X-ray diffraction patterns of the alloys containing 46, 50, 55. 60, and 62 atomic pct Al were identical, except for a slight shift in the positions of the reflections. The alloy containing 55 atomic pct Al was chosen far structure determination. The powder pattern contained 37 reflections and the a, doublets were well resolved in the back-reflection range. The relatively small number of reflections and the sharpness of the pattern were considered as an indication of a simple atomic arrangement in a relatively small unit cell of high symmetry, and an attempt was therefore made to solve the structure without single crystal work. A satisfactory fit was found on a large scale Hull-Davey tetragonal chart for an axial ratio of approximately 1.45. All the reflections were readily indexed and the tentative unit cell dimensions were computed to be a - 2.81 kX, c - 4.07 kX, and c/a = 1.448. The indices (hkl) and the values of d and sin2 A are given in Table I. The agreement between observed and computed sin2 is quite satisfactory. Once the unit cell was known, the details of the structure were readily found. First, it is obvious that the number of atoms per unit cell cannot exceed two, since more than two atoms would lead to a density greater than that of pure titanium. The next logical step is to assume that the two atoms occupy the corner and center of the cell. If this assumption is correct, the structure may be described as follows: space group D1, - P 4/mmm; one titanium atom in a: 000; one aluminum atom in d: 1/2 1/2 1/2: and no extinctions. Assuming this structure to be the correct one. intensities were computed by means of the usual equation: sin2 0 cos 0 where F is the structure factor; 8, the Bragg angle. and p, the multiplicity factor. The calculated values of intensities are compared in Table I with the visually estimated intensities. The agreement is quite satisfactory and the structure is therefore confirmed. The crystal structure of TiAl just described is the same as that of AuCu ordered. The AuCu ordered

Jan 1, 1953

COLORADO SCHOOL OF MINES On February 21, 1918, the students of the Colorado School of Mines who are members or have applied for membership in the Institute re-organized the "Scientific Society of The Colorado School of Mines,? which has for its .purpose the study and discussion of mining problems. Its meetings will be held on the first and third Wednesdays of each month, in Guggenheim Hall. Following are the officers R. W. GIBSON, President, W. A. CONLEY, Vice-president, F. L. SERVISS, Sec'y.-Treas. At present there are 62 members, seven of whom are Faculty members. F. L. SERVISS, Secretary.

Jan 5, 1918

(Members are urged to send in for this column any notes of interest concerning themselves or their fellow-members.) Members and guests who registered at Institute headquarters during the period May 10 to June 10: Rene de Sallier, New York, N. Y. Karl Koelle, San Luis Potosi, Mexico. H. S. Chamberlain, Chattanooga, Tenn. W. A. Wilson, Salt Lake City, Utah. F. R. Pretyman, Oyster Bay, N. Y. P. E. Barbour, Candor, N. C. R. Eustice, South Australia. R. C. Blanchard, London, England. F. C. Darby, South Australia. S. K. Tong, Shanghai, China. Grant H. Tod, Livermore, Cal. W. T. Burns, Great Falls, Mont. J. Murray Clark, Toronto, Canada. F. W. Bacorn, Butte, Mont. J. H. Adams, Birmingham, Ala. H. B. Pulsifer, Chicago, 111. A. Faison Dixon, Encontrador, Venezuela. Dwight E. Woodbridge, Duluth, Minn. C. W. Merrill, San Francisco, Cal. Heinrich Ries, Ithaca, N. Y. Francis Church Lincoln, Boston, Mass. H. L. Slosson, San Francisco, Cal. Frank R. Corwin has severed his connection with the Arizona Copper Co., Clifton, Ariz., and is now shift superintendent at the smelting plant of the Copper Queen Consolidated Mining Co.. Douglas. Ariz.

Jan 7, 1914

By O. L. Culberson, J. J. McKetta

Experimental and smoothed data are presented for the solubility of methane in water for temperatures of 77, 100, 160. 220. 280, and 340°F at prejsures to 10.000 psia. The minimum solubility phenomenon exists within this temperature range to pressures of 10.000 psia. INTRODUCTION The study of hydrocarhon-water systems has been extended to the determination of the solubility of methane in water at pressures to 10.000 psia. The experimental apparatus, and the technique. are the same as that previously reported.' - In the determination of the solubility of methane in water, only two phases were present at temperatures between 100 and 340°F. these being the vapor phase and the water-rich. liquid phase. The determinations involved the agitation of methane and water in an equilibrium cell. withdrawal of a sample from the liquid phase, and the analysis of the sample. The experimental value; determined in this manner are pre9ented in Table I and in Fig. 1. Since experimental and analytical techniques have improved since the initial experiments with this equipment, it was decided to repeat the 77°F isotherm determinations.' Fig. I shows that the individual results for this isotherm are much

Jan 1, 1951

By J. J. McKetta, O. L. Culberson

Experimental and smoothed data are presented for the solubility of methane in water for temperatures of 77, 100, 160. 220. 280, and 340°F at prejsures to 10.000 psia. The minimum solubility phenomenon exists within this temperature range to pressures of 10.000 psia. INTRODUCTION The study of hydrocarhon-water systems has been extended to the determination of the solubility of methane in water at pressures to 10.000 psia. The experimental apparatus, and the technique. are the same as that previously reported.' - In the determination of the solubility of methane in water, only two phases were present at temperatures between 100 and 340°F. these being the vapor phase and the water-rich. liquid phase. The determinations involved the agitation of methane and water in an equilibrium cell. withdrawal of a sample from the liquid phase, and the analysis of the sample. The experimental value; determined in this manner are pre9ented in Table I and in Fig. 1. Since experimental and analytical techniques have improved since the initial experiments with this equipment, it was decided to repeat the 77°F isotherm determinations.' Fig. I shows that the individual results for this isotherm are much

Jan 1, 1951

By A. Lee Barrett

MINE power systems are quite different in many respects from those usually found in industrial plants. Wide areas are served, usually by a circuit which is connected continuously throughout the mine. In some mines one circuit may run for several miles. There are usually a small number of branch circuits involved considering the areas covered. In all mines circuits are sectionalized at approximately 2500 ft by means of switches. All, or at least most branch circuits, are sectionalized by manual switches. In some mines automatic sectionalizing gear is used to give automatic overload protection at various strategic points in the system. However at the present time this practice is by no means general. Circuit hazards in mine electric systems are much greater than those encountered in industrial installation. Falls of roof or coal form a prominent hazard which may break the wire and may cause ground faults or partial ground faults. Ordinarily electric cables in mines cannot be insulated, there- fore n fall can easily cause a short circuit. The humidity in most mines is greater than 95 pct and at many points the floor is wet and water may be dripping from the roof over insulators, increasing the difficulties at that point. All electrical circuits must be contained in the mine entries and space must also be provided for moving equipment and men. Consequently, because of the proximity of the power circuit to the locomotives and other moving equipment a considerable hazard is introduced. Mine haulage track is not always good so that wrecks are not unusual. A wreck frequently does consider- able damage to the power circuit. A ground fault frequently results in heavy arcing which may be sustained. If the damage to the circuit is mechanical it may mean that a trip of coal has wrecked and a great deal of coal dust could be thrown into the air as a result of the wreck. The arc resulting from the ground fault could ignite the coal dust and a mine explosion would result. A continued ground fault, even if high in resistance, also may ignite the coal which may be on the bottom of the entry and a mine fire may result. The fire hazard is one of the most serious hazards in any coal mine. Since the mine circuit is completely interconnected, any power interruption resulting from a fault seriously affects the whole operation. Unless the mine is completely sectionalized the whole mine will be shut down until the fault is located and the proper switches are opened to isolate the fault; even then large areas may still be without power if they are supplied by a single power circuit which may be opened in isolating the fault. Further, mine power is usually supplied from a relatively small number of substations and the fault may be on a stub feed circuit, in which case power cannot be restored until the fault has been cleared. Practice varies widely with respect to placement of substations. Usually they will be spaced from I to 2 miles in 500V operations, and at 1 mile intervals or less, in 250 V operation. Actual mining operations govern substation placements and no general rules can be followed.

Jan 1, 1948

By A. Lee Barrett

Mine power systems are quite different in many respects from those usually found in industrial plants. Wide areas are served, usually by a circuit which is connected continuously throughout the mine. In some mines one circuit may run for several miles. There are usually a small number of branch circuits involved considering the areas covered. In all mines circuits are sectional-ized at approximately 2500 ft by means of switches. All, or at least most branch .circuits, are sectionalized by manual switches. In some mines automatic sectionalizing gear is used to give automatic overload protection at various strategic points in the system. However at the present time this practice is by no means general. Circuit hazards in mine electric systems are much greater than those encountered in industrial installation. Falls of roof or coal form a prominent hazard which may break the wire and may cause ground faults or partial ground faults. Ordinarily electric cables in mines cannot be insulated, therefore a fall can easily cause a short circuit. The humidity in most mines is greater than 95 pct and at many points the floor is wet and water may be dripping from the roof over insulators, increasing the difficulties at that point. All electrical circuits must be contained in the mine entries and space must also be provided for moving equipment and men. Consequently, because of the proximity of the power circuit to the locomotives and other moving equipment a considerable hazard is introduced. Mine haulage track is not always good so that wrecks are not unusual. A wreck frequently does considerable damage to the power circuit. A ground fault frequently results in heavy arcing which may be sustained. If the damage to the circuit is mechanical it may mean that a trip of coal has wrecked and a great deal of coal dust could be thrown into the air as a result of the wreck. The arc resulting from the ground fault could ignite the coal dust and a mine explosion would result. A continued ground fault, even if high in resistance, also may ignite the coal which may be on the bottom of the entry and a mine fire may result. The fire hazard is one of the most serious hazards in any coal mine. Since the mine circuit is completely interconnected, any power interruption resulting. from a fault seriously affects the whole operation. Unless the mine is completely sectionalized the whole mine will be shut down until the fault is located and the proper switches are opened to isolate the fault; even then large areas may still be without power if they are supplied by a single power circuit which may be opened in isolating the fault. Further, mine power is usually supplied from a relatively small number of substations and the fault may be on a stub feed circuit, in which case power cannot be restored until the fault has been cleared. Practice varies widely with respect to placement of substations. Usually they will be spaced from I to 2 miles in 500V opera-tions, and at I mile intervals or less, in 250V operation. Actual mining operations govern substation placements and no general rules can be followed.

Jan 1, 1949

By J. B. Austin

In considering so complex a process as the smelting of iron in the blast furnace, there is obviously no single method of calculating efficiency that gives a complete appraisal of the performance of the furnace in all its several functions; such a comprehensive view as is required to evaluate producing ability and manufacturing costs for a given furnace is obtained only by considering a group of efficiencies, each of which measures the performance of the furnace from a specific point of view. For instance, there is the rate of consumption of coke per ton of iron produced, which is probably the most frequently considered, and is certainly one of the most important of the group; there is the efficiency of recovery of iron charged; and there is the efficiency of utilization of energy, both chemical and thermal, in some respects related to the former two yet in other ways independent of either. In each of these cases, the calculation can be made on the basis of the actual input of material or energy, or it may be based upon a comparison with the minimum amount of material or energy that would be required by a "perfect" furnace; that is, an imaginary ideal furnace analogous to the perfect steam engine used in thermodynamic calculations, when this furnace is operating under optimum conditions. Moreover, the furnace itself can be considered either as an apparatus for smelting iron or, taking a somewhat less common point of view, as a gas producer yielding iron and slag as by-products. Some of these efficiencies are more useful than others, yet all are instructive; for a knowledge of them not only leads to good engineering practice from the standpoint of control but also discloses limits to possible improvement in performance and indicates the lines along which improvement is most likely to be profitable. Estimates of some of these efficiencies have appeared from time to time in the literature, but many of them are open to objection because they are based on data of uncertain accuracy or because they have been calculated without due regard to the thermodynamic principles involved. Moreover, so far as the author is aware, no general survey of the blast furnace comparing its performance in all the aspects mentioned has been made.

Jan 1, 1938

By S. S. Nosseir

INTRODUCTION General Quebec Cartier Mining Co. (QCM) , Canadian subsidiary of U . S. Steel Corporation, operates two iron ore concentrators at the North Shore of Quebec, Canada. In its own operation at Mount-Wright, over 45 million tons per year of crude ore are ground, fully autogenous in six 32' x 11-1/2 Dominion Engineering mills. At Fire Lake/Lac Jeannine operation, property of Sidbec Normines Incorporated (SNI), over 11 million tons per year of crude ore are now ground, fully autogenous, in twelve 18' x 5' Hardinge mills. The Lac Jeannine concentrator was originally constructed by Q.C.M. in 1960 and. was since operated, in beneficiating the Lac Jeannine ore deposit until fully depleted in 1977, at an annual capacity of 18 million tons of crude per year. In 1976, the new company, Sidbec Normines, was incorporated by Sidbec Corporation, British Steel Corporation and Quebec Cartier Mining Co., in proportional holding of approximately 51 to 41 to 8 percent respectively. Since the acquisition of the Fire Lake ore deposit and Jeannine concentrator by the new company, QCM has continued managing the mine and concentrator operaions for S idbec Normines Inc . Flow-Sheet Design Jeannine concentrator comprises twelve identical mill lines grinding the ROM from 8" top size to 14 mesh nominal size range. Driven by 600 H.P. motor at 13.4 RPM corresponding to 73% critical speed, each mill is operated in a closed circuit with two 5' x 12' vibrating screens dressed with 2.5 x 25 mm slot opening cloth. The ground product or screen undersize at an average grade of 33.4% Fe is beneficiated by gravity separation to produce 66.5% Fe concentrate grade at 42.7% weight recovery. At the rougher stage, each mill line con-tains 128 Humphrey 13- 1/2" pitch spirals which produce a final tailing, middling and concentrate products. The rougher concentrate - so produced - is classified by 64 Humphrey pocket sizers into an underflow, which is a final concentrate product and an overflow which is further upgraded by 64 cleaner spirals. The cleaner concentrate together with the pocket sizer 's underflow are dewatered on a 13'[6] horizontal filter. The rougher middling together with the cleaner tailing are recycled in a closed circuit with the spirals. The rougher tailing after dewatering by hydrocyclones is disposed of by a

Jan 1, 1982

By J. B. Austin

In considering so complex a process as the smelting of iron in the blast furnace, there is obviously no single method of calculating efficiency that gives a complete appraisal of the performance of the furnace in all its several functions; such a comprehensive view as is required to evaluate producing ability and manufacturing costs for a given furnace is obtained only by considering a group of efficiencies, each of which measures the performance of the furnace from a specific point of view. For instance, there is the rate of consumption of coke per ton of iron produced, which is probably the most frequently considered, and is certainly one of the most important of the group; there is the efficiency of recovery of iron charged; and there is the efficiency of utilization of energy, both chemical and thermal, in some respects related to the former two yet in other ways independent of either. In each of these cases, the calculation can be made on the basis of the actual input of material or energy, or it may be based upon a comparison with the minimum amount of material or energy that would be required by a "perfect" furnace; that is, an imaginary ideal furnace analogous to the perfect steam engine used in thermodynamic calculations, when this furnace is operating under optimum conditions. Moreover, the furnace itself can be considered either as an apparatus for smelting iron or, taking a somewhat less common point of view, as a gas producer yielding iron and slag as by-products. Some of these efficiencies are more useful than others, yet all are instructive; for a knowledge of them not only leads to good engineering practice from the standpoint of control but also discloses limits to possible improvement in performance and indicates the lines along which improvement is most likely to be profitable. Estimates of some of these efficiencies have appeared from time to time in the literature, but many of them are open to objection because they are based on data of uncertain accuracy or because they have been caleulated without due regard to the thermodynamic principles involved. Moreover, so far as the author is aware. no general survey of the blast furnace comparing its performance in all the aspects mentioned has been made.

Jan 1, 1938

By J. B. Austin

In considering so complex a process as the smelting of iron in the blast furnace, there is obviously no single method of calculating efficiency that gives a complete appraisal of the performance of the furnace in all its several functions; such a comprehensive view as is required to evaluate producing ability and manufacturing costs for a given furnace is obtained only by considering a group of efficiencies, each of which measures the performance of the furnace from a specific point of view. For instance, there is the rate of consumption of coke per ton of iron produced, which is probably the most frequently considered, and is certainly one of the most important of the group; there is the efficiency of recovery of iron charged; and there is the efficiency of utilization of energy, both chemical and thermal, in some respects related to the former two yet in other ways independent of either. In each of these cases, the calculation can be made on the basis of the actual input of material or energy, or it may be based upon a comparison with the minimum amount of material or energy that would be required by a "perfect" furnace; that is, an imaginary ideal furnace analogous to the perfect steam engine used in thermodynamic calculations, when this furnace is operating under optimum conditions. Moreover, the furnace itself can be considered either as an apparatus for smelting iron or, taking a somewhat less common point of view, as a gas producer yielding iron and slag as by-products. Some of these efficiencies are more useful than others, yet all are instructive; for a knowledge of them not only leads to good engineering practice from the standpoint of control but also discloses limits to possible improvement in performance and indicates the lines along which improvement is most likely to be profitable. Estimates of some of these efficiencies have appeared from time to time in the literature, but many of them are open to objection because they are based on data of uncertain accuracy or because they have been caleulated without due regard to the thermodynamic principles involved. Moreover, so far as the author is aware. no general survey of the blast furnace comparing its performance in all the aspects mentioned has been made.

Jan 1, 1938

By Raymond M. Howe

Some time ago, a prominent engineer asked a representative of the firebrick industry to prepare a comprehensive paper on blast-furnace refractories. It was to have been the purpose of this paper to gather practical experiences from widely different sources, in the hope of determining what kind or kinds of firebrick were best adapted for certain uses. This engineer sent questionnaires to the large consumers of blast-furrace linings and stovebrick and received very detailed replies as to their experiences. These replies are conflicting, but, having been turned over to the writer, form the basis of this attempt to meet the request. Because of the impossibility of making final recommendations at this time, a few of the more interesting and commonly known methods of studying refractories are also discussed, as well as the more recent developments in this field. The first reply to the questionnaire stated that good service was being secured from the linings. This satisfactory condition was attributed to the use of a very good distributor. The writer believed that manufacturers try to and do produce a good product and that the proper handling of equipment is of the utmost importance, if good results are to be secured. The type of bricks used was not mentioned. The second reply stated that good life was secured from the furnace linings, but that the writer had had soft-fired stovebrick crumble and, for that reason, preferred a hard-fired product for use in this position. The type of bricks used was not mentioned. The third reply stated that firebrick had crumbled from the mantle up; the cause was not known. The fourth reply stated that the stove linings were satisfactory. Hand-made furnace linings failed after.4 years' service so a steam-pressed lining- was installed which was still good after 4 years of service. Consequently, a second steam-pressed lining was installed but this disintegrated at the top after the furnace had been in blast 7 months. The firebrick manufacturer attributed this failure to the expansion of iron blocks that had been installed at the stock line to prevent erosion. The fifth reply stated that the present "hard-burned" hand-made furnace lining had worn away very uniformly to the depth of only 5 in.

Jan 1, 1920

By M. E. Lombardi

THE Coalinga oil field is located on the west side of the San Joaquin Valley, California. The structure is in general a monocline, the edges of the oil horizon resting on the foot hills and dipping gently toward the east. One prominent anticline occurs plunging southeast. The earlier drilling was done in the foot hills comparatively near the outcrop and the wells were shallow. The sands were followed eastward and, in the case of the anticline, along the plunge, the wells becoming deeper and deeper until the depth of 4,000 ft. was reached and passed. There is nothing to show that the oil will not be found in quantity at still greater depth. In fact, some of the best producers have tapped the sand at close to 4,000 ft. The recovery of oil still farther to the east, and therefore at greater depth, seems to be mainly a question of drilling. In this territory the formations drilled through are chiefly sands and shales; they will not "stand up" in an open drilling hole; the casing has to be carried close to the bit, and it is always difficult to keep the casing free for any considerable distance. Ability to carry casing of comparatively large diameter without conductor pipes for distances of 2,000 or 3,000 ft. or over is desirable in such territory chiefly for two reasons. It makes it possible to enter the oil sand with a pipe of ample diameter; it eliminates one or more expensive strings of casing which act only as conductors for the water string, and furthermore, in territory where waters are encountered which corrode steel rapidly, it makes possible the construction of a rust- and alkali-resisting water string. It is always desirable to shut off top waters, which may lie within 100 ft. or less of the oil sand, with 10-in. pipe. Where the depth is so great that a practical weight of 10-in. pipe will not withstand the probable collapsing pressures, 8 ¼ in. at least is desirable. About the limit of rotary drilling to date in California seems to be the setting of the 10-in. string at 3,200 ft., although the rapid advance in rotary work during the past year seems to indicate that this depth may soon be increased. It is my purpose now, however, to treat only of cable-tool drilling.

Jan 2, 1915

By H. W. Winkler, Discussion, V. L. Forsyth

The development of a new gar lift valve has removed many of the obstacles limiting over-all gas lift efficiency. The valve is pressure charged in place in a well, and operating pressure can be changed without pulling. Temperature and gas gradient compensation have been eliminated. Intermitting and constant flow installations, using conventional pressure-charged valves, are compared with designs incorporating the new automatic stabilizer controlled valve. Specific well performance data are presented. The means of obtaining and controlling multiple-point injection of gas are explained and contrasted with single point injection. The effect on small diameter strings and multiple string installations is discussed. The effect of flowing temperature gradient on a design using conventional pressure-charged valves, the limitations imposed by such temperature, and the obvious benefits which result from the use of automatic stabilizer controlled valves are shown. Reduction of gas requirements is stated mathematically and demonstrated with specific examples which are identified as to oil company and well. Increased fluid recovery resulting from greater drawdown is illustrated. The economic advantages are mentioned with emphasis on reduction of capital expenditures, as well as reduction in operating expenses. INTRODUCTION Much of the engineering and research work performed with gas lift has been based upon the supposition that under producing conditions gas would be injected at a single point in a well. Many tests were performed in determining the optimum size of the opening through which gas would be passed and the exact placement of the valve equipment. The objectives in a gas lift installation are to unload the well and to efficiently produce the well after it is unloaded. Until recently, the function of upper valves was simply to unload the well. An operating valve located at a proper place in a well served as a single injection point for lifting liquid to the surface. The basic problem in using such equipment is to arrive at operating depth with sufficient operating pressure to efficiently lift the liquid. Few gas lift installations are designed with all required well data available. However, regardless of the quality and availability of data, and regardless of the accuracy of the design, there is the practical problem of preparing and placing gas lift equipment in the well in exact accordance with the plans and intentions of the engineer. In recognition of these and many other problems, attention was directed toward a means of removing the need for temperature and gas gradient compensation, a way to change the gas lift installation to meet the changing requirements of the well itself, and a manner of compensating for changes in oil-water content, fluid volumes, PI, and bottom-hole pressure resulting from flood or repres-suring activity. A device which could actually set and change the operating pressure of gas lift valves in the well would meet most of these needs. There now exists a new gas lift valve which fulfills the original objectives and in so doing actually accomplishes much more. With this device, called an Automatic Stabilizer Controlled valve, there is an exact matching of valve operating pressure and well conditions. Gas is injected at more than one point in the well, thereby making controlled multiple-point injection of gas a reality. Installation calculations are materially simplified and the need for basic temperature and gas gradient data no longer exists. Section I—The Automatic Stabilizer Controlled Valve The operation of the ASC valve is more easily understood by first studying a conventional pressure-charged valve and then inserting the stabilizer element to observe the changes which result. Fig. 1 is a schematic of a conventional precharged pressure-operated valve. The pressure in the dome pd, acts downward, against the area of the bellows A,, to hold the valve closed. The resulting force Bf plus the spring effect of the bellows Sf represents the total force acting to close the valve. Correspondingly, the tubing pressure ptbg works upward against the port area Atbg and the casing pressure pc works upward against a portion of the area of the bellows

Jan 1, 1965

By M. F. Williams, L. G. Hendrickson

IN classification of pulps bearing magnetized ferromagnetic particles, depolarizing is of great importance. If size separation is to be effective, particles must be individual rather than in floes. Depolarizing is also practiced in heavy medium separations in which ferrosilicon or magnetite is the medium. When particles of ferromagnetic material have been removed from a magnetic field, residual magnetism causes agglomeration. The term depolarizing refers to the operation of reducing or eliminating this residual magnetism and may thus be considered magnetic deflocculation. The terms demagnetizing and randomizing are also used. At the Research Laboratory of Oliver Iron Mining Div. in Duluth a method was developed for measuring depolarization of the pulp of ferromagnetic material. Experiments were made with Mesabi taconite,' a natural magnetite of low coercive force. Ferromagnetic materials of higher coercive force, such as lodestone or the artificial magnetite produced by reduction roasting of hematite, present a more difficult problem, which was not within the scope of this investigation. It is possible, however, that some of the techniques evolved for measuring and calculating electrical characteristics of alternating current coils would be of use in depolarizing high coercive force material, particularly in conjunction with high-frequency alternating current, as proposed by Hartig and others.'," Properties of Ferromagnetic Materials:'7 Experimental work, described below, has shown that if a sample in the magnetized state is heated above the Curie point and cooled, much of the preferred orientation is destroyed and the sample is substantially depolarized. It has been thought that when the sample cools below the Curie point the domains cancel each other, leaving a zero net moment. However, such particles still exhibit a tendency to cohere, and undoubtedly this is caused by the forces of residual magnetism. -AS measured by the percent depolarization, this tendency is reproducible for any sample upon repeated heating above the Curie point and subsequent cooling and is independent of the initial state of magnetization. It is postulated, therefore, that as the material is cooled below the Curie point the domains in any particle do not completely cancel each other, but rather are preferentially oriented to some extent. Mechanism of Depolarizing with Alternating Current Magnetic Fields: It is believed that when ferromagnetic material is passed through an alternating magnetic field, depolarizing occurs in the decaying portion of the field. As the particles pass through the portion of highest intensity they become magne- tized. If the particles are not free to move, the polarities of the particles will be reversed (by a mechanism similar to that described above for magnetism) at a frequency equal to that of the applied field. As the material moves through the decaying field, intensity levels become such that a domain does not completely reverse, but stops on an axis of easy magnetization. By the time the material reaches the point of zero field intensity, a state of fairly random orientation of domains is achieved. If conditions are such as to give a completely random orientation, the particle will have little or no external magnetic field, and a pulp of such particles will be depolarized. Previous Work In 1918 E. W. Davis was granted a patent for demagnetization of magnetite pulps.' His method consisted of passing the pulp through a tapered coil, activated by alternating current of normal frequency (60 cycle). This method, with minor modifications, has been used almost universally in all pilot plants and commercial installations in which depolarization of low coercive force materials has been required. Hartig, Onstad and Foot2,3 avd made a detailed study of the factors involved in depolarizing both low (below 100 oersteds) and high (above 100 oersteds) coercive force material. They developed a method for evaluating the relative degree of depolarization of any pulp based on the settling characteristics of the pulp. Their standard of comparison was a sample heated to above the Curie point and cooled in a zero field, all in a neutral atmosphere.* For low coercive force material they found treatment.Thisprocedure is subsequently called, in this report,theCurie that results equivalent to Curie treatment could be

Jan 1, 1957

By Jack White

MUFULIRA Copper Mines, Limited, conducted tests on the concentrator in 1942 to determine which type of ball mill, that is, Conical Trunnion overflow or Cylindrical Grate discharge, produced the greater quantity of finished product at the lowest cost. All the data which follows applies to Mufulira ore and to the particular degree of grinding produced with the classifiers installed at Mufulira concentrator. JACK WHITE, Member AIME, is Concentrator Superzntendent, Mufulira Copper Mines, Ltd., Mufulira, Nortltern Rhodesia. Sun Francisco Meeting, February 1949. TP 2729 B. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Feb. 28, 1949. Test Procedure: The test was carried out using ball mills No. 1 to 9 (conical trunnion overflow) as the standard with which to compare No. 10 (cylindrical grate discharge mill). The conical trunnion overflow ball mills are 10 ft diam by 72 in., and the cylindrical grate discharge mill is a No. 99. Each mill runs in closed circuit with a Dorr type DSFX classifier 8 ft wide by 30 ft long. The feed is considered as being identical to all ten units. The figures for the test were derived as follows. Dry Tons Milled: For No. 10 unit (cylindrical grate discharge mill) this figure was taken from the ball mill weightometer, which was corrected for error, and then for moisture content. The weightorneter was checked twice a week during the test, and the error did not exceed 3 pct at any time. The moisture was obtained from a 24 hr moisture sample taken by the ball mill operator hourly, and placed in a closed bucket. The sample was cut where the conveyor drops the feed into the ball mill chute. The Mill Sampling Department made a daily moisture determination on this sample. To obtain the dry tons on units No. 1 to 9 the No. 10 figure was deducted from the month's dry tons milled figure, which, in turn, was taken from the crusher weightometer tons corrected for weightometer error and moisture. All tons reported in this test are short tons of 2000 lb. Ball Mill Hours: This was taken daily from the Shift Boss reports giving actual running hours for each unit. Screen Analyses: Coarse Heads (Ball Mill Feed): The monthly composite coarse heads screen analysis, done by the Mill Sampling Department, was used for the minus 200 mesh in the feed to the mills. F. X. Classifier Overflow: The ball mill operators took an hourly 24 hr sample of the overflows of units No. 1 to 9. This was a hand sample, but cut with a special cutter, and is believed to be very reliable. On No. 10 unit, an hourly 24 hr sample, and an hourly shift sample were cut in the same way. These five samples were split, filtered, and dried by the Research Department and were then wet screened on a 200 mesh screen. The average of the three shifts samples on No. 10 unit had to check the 24 hr sample within 1.5 pct to be acceptable. A weighted composite was cut each day from both 24 hr samples, and at the end of each month it was screened as follows: Wet screened through 325 mesh, the plus 325 then screened on a rotap for 15 min through 65, 100, 150, 200, and 325. This was done in duplicate. All splitting, drying, and screening of samples was done by one of the Mill Research staff, to insure accuracy. F. X. Classifier Overflow Densities: These were taken from the Shift Boss reports. Densities were taken by the ball mill operator every hour. Circulating Loads: These were taken a number of times during the test both by direct method (weighing a measured portion of classifier rake sand, and deducting moisture) and by calculation using the formula:

Jan 1, 1951