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Aamot, Olav Crone, (M'29) Chem. Engr., Elektro¬kemisk, Raadhusgt, 23; Oslo, Norway. Abadilla. Quirico A., (M'33) Min. Engr., Dir., Bureau of Mines, Manila, P. I. Abbott, Agatin T., ( J'40) Asst. Geol., Trench Mine, Patagonia, Ariz. Abbott, Clarence E., (M'04) Civil Engr., V.P., Tennessee Coal Iron & R. R. Co.. 159'3 Brown Marx Bldg.. Birmingham, Ala. Abel, Raymond L., (M'28) Chem. Engr., Chief of Inspection Div., Pittsburgh Chem. Warfare Pro¬curement Dist., American Batik Bldg., Pitts¬burgh. Pa. Abel, Walter D., (M'32) Min. Engr.. Div. of Cor¬porations, State of Calif.. 403 California State Bldg., Los Angeles, Calif. Abele. Louis T., (M'31) 4800 Patterson Ave.. Richmond, Va. Abell, Leo M., (M'311 Research Met., Mount Mor¬gan Ltd., Mount Morgan, Queensland, Australia. Abenius, Hakan V., (M'37) c/o Falu Gruva, Falun, Sweden. Abernathy. G. E., (M '33S) Geol., Kansas State Geol. Survey. Pittsburg, Kan. Abernathy, Jack. H.. (M'40) Chief Engr., Sunray Oil Co.. Philtower B1dg., Tulsa. Okla. Ahorn, Robert H., (M'19) Asst. Met., Research Lab., U. S. Steel Corp., Lincoln Highway, Kearny, N. J. Abraham, A. W., (M'30) Petr. Engr., Richfield Oil Corp., 555 S. Flower St., Los Angeles. Calif. Abramowitz, William L., (J'40) Chem., National Oil Products Co., Harrison, N. J. Abrams, Charles Jackson, (M'36) Gen'l Supt., Climax Molybdenum Co.; Climax, Colo. Abstein, Henry T., (M'27) Yellow Pine, Idaho. Ackerman, D. E., (M'35) McKinsey & Co., 2 Wall St., New York, N. Y. Ackerman. E. Dumont, (A'3S) Petr. Geol.. Dist. Geol., Quiriquire' Dist., Standard Oil Co. of Venezuela, Caripito, Venezuela.

Jan 1, 1941

By Harry Klemic

Zirconium and hafnium minerals are used industrially both as minerals valuable for their chemical and physical characteristics and as ores of zirconium and hafnium. The principal zirconium-hafnium-bearing minerals, their composition, and their ranges in content of zirconium and hafnium dioxides are listed in Table 1. The supply of zirconium and hafnium is based almost entirely upon the supply of the minerals zircon and baddeleyite. Eudialyte may eventually also be of importance as a source of these elements. Zircon, which generally contains nearly 49% zirconium and between 0.4 and 1.5% hafnium, is the principal industrial mineral of this group. Baddeleyite, a commercially important but less abundant mineral, contains as much as 73% zirconium and 0.4 to 1.7% hafnium. Eudialyte has a wide range in content of these elements, but analyses of material from several localities indicate values of 6 to 11 % zirconium and 0.07 to 0.4% hafnium. Zircon is a mineral consisting principally of zirconium, silicon, and oxygen that crystallizes from magmas during the formation of igneous rocks. Zircon also occurs in veins and in metamorphic rocks. It is in the tetragonal crystal system. It commonly forms as small well-developed prisms having pyramidal terminations but also occurs as irregular-shaped grains. Zircon has a hardness of 7.5 on the Mohs' scale. Its specific gravity is commonly between 4.68 and 4.70 (Dana, 1932), but the specific gravity varies somewhat depending upon com- position and state of alteration. Zircon has no good cleavage directions and is brittle, breaking with a conchoidal fracture. Zircon of the composition ZrSiO, would contain 67.2% Zr02 and 32.8% SiO2. However, in addition to about 1 % of hafnium, minor amounts of thorium, uranium, rare earth elements, yttrium, calcium, magnesium, iron, aluminum, phosphorus, and hydrogen, and trace amounts of other elements are commonly present in the mineral (Frondel, 1957). Radiogenic lead and other daughter products from the radioactive decay of uranium and thorium also accrue in the mineral. Radiation damage to the structure of zircons that contain appreciable amounts of uranium and thorium may alter the internal structure of the crystals to

Jan 1, 1975

By R. W. Raymond

Edward Cooper was born in New York City, Oct. 26, 1824. His father, Peter Cooper, to say nothing of manifold reasoils for fame as an inventor and philanthropist, deserves to be remembered as a pioneer in the establishment .of public schools ; and it is not surprising that. he sent his son to one of the democratic institutions in which he was himself so ardently interested. From the public school Edward Cooper went to Columbia College, where he made the acquaintance of Abram S. Hewitt, who was first his tutor, afterwards his traveling companion, brother-in-law, business partner and life-long friend.' For twenty years I was the consulting engineer of the firm of Cooper & Hewitt, and also connected with the work of the Cooper Union, in which they were both deeply interested and faithfully active. On the basis of the intimate knowledge thus gained, I venture an estimate of Edward Cooper's character and career which, if not fully comprehensive, may facilitate a juster estimate than would be reached by the student of those data which the modesty of the man permitted to be placed upon public record.

Jan 1, 1907

By D. T. Peters, C. R. Cupp

Aging of commercial 18 pct Ni maraging steel and of related binary, ternary, and quaternary alloys has been studied by electrical resistivity, X-ray diffrac-tion, and hardness measurements. The aging process is in three stages: 1) recovery of martensite; 2) two precipitation reactions; 3) formation of nickel- rich austenite. The main objective of this work was to discover the nature of the Co-Mu interaction. Addition of cobalt to the Fe-18 pct Ni-5pct Mo alloy results in acceleration of the pvecipitation reaction by increasing the super-sat~lration 0f molybdenum. The cobalt-containing al- loys show no incubation period for the precipitation reaction, but the alloy without cobalt shows an initial slow reaction. These results are interpreted in terms of precipitation on dislocations. At low aging temperatures, the presence of cobalt leads to the appearance of a second type of precipitate which is a particularly effective hardener. Cobalt was found to retard the formation of austenite. Overaging is a combination of classical overaging and the formation of austenite in the structure. he- activation energy for the precipitation reaction determined from resistivity ,measurements is 42.8 kcal per mole and the growth of the precipitate is found to obey at law at high tenzperatures and a tU3 law at low temperatures. HE mechanism of strengthening of the 18 pct Ni maraging steels has been studied extensively since 1962.'-? The strengthening is due to precipitation hardening of the martensitic matrix. The main precipitate is an inter metallic compound involving molybdenum, but there is disagreement about its identity.

Jan 1, 1967

By Lyon F. Terry

IN contemplating the prospects of natural gas being transported from the fields where it is produced to such distant points as Philadelphia, New Jersey, New York City, and New England, let us review the advance that natural gas has made and the projects currently scheduled for construction. This may give us a perspective of the point in the long-range development of the natural gas business in which we now find ourselves, and some basis for judging the further expansion of the industry. Authentic statistics of the total marketed production of natural gas in the United States have been compiled by the Bureau of Mines since the year 1906. The growth of the business is indicated by the following data on marketed production for each ten-year interval since 1906:

Jan 1, 1947

By R. W. Diamond

THREE mills are now operated by the Consolidated Mining & Smelting Co. of Canada, Ltd:: (1) The Sullivan concentrator, Chapman, B. C. (near Kimberley, B. C.), (2) the St. Eugene concentrator, Moyie, B. C., and (3) the Tadanac (customs) concen-trator, Trail, B. C. Differential flotation is the only concentration method employed in all three plants. Tables are used for pilot work only. Flotation equip-ment and control are similar in the three mills, and the flow sheets at the St. Eugene and Tadanac mills, while much simpler, are modeled after that of the Sul-livan concentrator. Successful differential flotation of Sullivan ore com-menced in the spring of 1920, in the pilot concentrator at Trail. During 1920, 193,804 tons were treated by this process. This operation continued without interrup-tion and with steady growth. For the first 7 months of . 1923, which concluded the pilot plant operation, ap-proximately 1100 tons were treated daily. I believe the Sunnyside plant of the United States Smelting &. Refining Co. was the only successful and large scale operation practising differential flotation prior to 1920. in North America. The Sullivan ore, because of its high iron content, was quite a different problem from that at Sunnyside. The entire tonnage of Sullivan ore had to be separated into three sulfide concentrate products-lead concentrate, zinc concentrate and iron concentrate. The choice and arrangement of equipment, and the original flow sheet for the Sullivan concentrator were controlled by the results of the exhaustive pilot mill test. These covered a continuous period of 3 1/2 years and involved the treatment of 1,076,368 tons of Sullivan ore. Satisfactory recoveries are made from feed; 11.0 per cent lead and 7.5 per cent zinc in the form of lead concentrate containing 69 per cent lead, 4.5 per cent zinc, and zinc concentrate containing 3.0 per cent lead and 49 per cent zinc. (The pure zinc mineral in Sulli-van ore is marmatite, 'containing approximately 54.5 per cent zinc and 10 per cent iron. It is apparent, there-fore, that 49 percent concentrate is high grade for this particular ore.)

Jan 8, 1927

By Edna A. Dancy, Ronald L. Pastorek, Gerhard J. Derge

In an earlier study1 in this laboratory, it was found that there is a minimum in the specific conductance of Ni-S melts in the region of the stoichiometric composition, Ni3S2. This and similar observations on other sulfide1 and telluride2 systems suggest that short-range order exists in the molten phase for several hundred degrees above the melting point, and a qualitative energy-band model has been proposed, with intrinsic semiconduction at the stoichiometric composition and degenerate electron and hole conduction at metal- and tellurium-rich compositions. This suggestion has been substantiated for the telluride systems by Seebeck coefficient measurements which showed a change in sign from negative on the metal-rich side to positive on the tellurium-rich side. In the present investigation, in order to substantiate still further the proposed model, Seebeck coefficient determinations have been extended to the Ni-S system over a composition range in the vicinity of Ni3S2. The experimental method was essentially the same as that used for the telluride systems, with minor modifications on cells for high sulfur melts to permit the use of tungsten wires and W— W-Re th~rmocouples instead of the chrome1 or alumel wires and chromel-alumel thermocouples originally used. Chrome1 and alumel both deteriorate very rapidly in atmospheres containing sulfur. The Ni-S alloys were prepared as described in the conductivity paper. Initially, a portion of the solidified melt was analyzed at the end of a run, but after this was found to be inaccurate because of segregation during cooling, the whole sample was analyzed. The nickel content, only, was determined, the sulfur content being obtained by difference. The results are summarized graphically in Fig. 1, and clearly show a change in the sign of the Seebeck coefficient in the region of the stoichiometric composition. This change of sign near Ni3S2, coupled with the minimum in the conductivity, is suggestive of an intrinsic semiconductor at the stoichiometric composition, with degenerate electron and hole conduction on either side of this composition, in keeping with the proposed model. The negative temperature coefficient of conductivity, normally indicative of metallic conduction, is not inconsistent with this model. In the completely degenerate region it would be expected, while in the vicinity of Ni3S2 the suggestion originally made by Argyriades, Derge, and pound3 for the Fe-S system can probably be applied. They proposed that the forbidden gap between the valence band and the conduction band is small, and that the increase in the number of the conducting species with increase in temperature is not large enough to compensate for the loss of mobility through thermal scattering. The research was sponsored by the Office of Naval Research under Contract NONR-760(22) Project NR 039-081. 'E. A. Dancy and G. J. Derge: Trans. Met. Soc. Aime, 1963, vol. 227, pp. 1034-38. 'E. A. Dancy: Trans. Vet. soc. AIME. 1964, vol. 233, pp. 270-77. 'D. Argyriades, G. Derge, and G. M. Pound: Trans. Met. Soc. . 1959, vol. 215, pp. 909-12.

Jan 1, 1965

By Cyril Smith

SEVERAL investigators, mainly concerned with the mechanical proper-ties of the alloys, have studied the so-called aluminum bronzes after various quenching and reheating treatments. Of these works, perhaps the most complete is that of Bouldoires,1 who studied changes of length, electrical conductivity, thermoelectric force, hardness and micro-structure oh tempering quenched alloys containing 5 to 20 per cent of aluminum, although the work of Carpenter and Edwards,2 Curry,3 Greenwood,4 Braesco,5 Matsuda;6,7 Grard8 and others has added much to our knowledge of the changes occurring during tempering. The so-called acicular ß decomposes on tempering at temperatures between 350° C. and the eutectoid, giving rise to a Widmanstätten structure of a needles in a matrix of d or fine a + d, associated with an increase in hardness and almost complete loss of ductility. Matsuda, studied the' eutectoid transformation by means of dilatation and electrical "conductivity measurements during cooling at different rates. He found that on cooling at the rate of 10° C. per minute there was a decrease in electrical resistance and in length instead of the usual increase which occurred with slow rates of cooling. On tempering the rapidly cooled samples, the increase commenced at about 400° C. and at 520° C. the resistance and length had reached the equilibrium values. These peculiarities he accounted for as being due to a transitional phase ß' being formed, ß having an electrical resistance and volume less than either ß or a + d.

Jan 1, 1933

By Ervin E. Underwood

IT has been recognized for many years that dis-persed particles have great value in raising the creep resistance of metallic alloys. In fact, some of the most successful high-temperature alloys owe their superiority to this factor above all others. A recent observation by Glen1," indicates that even better high-temperature resistance to deformation results in alloys capable of two successive precipitation reactions. The ramifications of this finding are of great interest. Unfortunately, our understanding of the reasons for such notable improvements has not kept pace with technical developments. In an effort to uncover some of the basic facts governing strengthening by particles, a program was initiated at the Battelle Memorial Institute with the primary goal of comparing creep strengths in single-phase and two-phase (precipitating) alloys. The effects of precipitation from solid solution on the creep properties of alloys are generally such that the resistance to creep is increased. However, if alloys are heat treated to different degrees of hardness before creep testing, it is often found that those possessing the maximum hardness initially do not necessarily have the maximum creep resistance. Depending on the stresses and temperatures involved other structures may prove more desirable. Considerations of the above nature indicate the importance of defining the initial state of the alloy. Here, it was decided to start with quenched single- phase alloys in which precipitation and creep deformation proceed concurrently. Equipment and Experimental Procedures Preparation of the Specimens—Four 25 lb ingots were prepared from 99.999 pct Cu and 99.996 pct Al to compositions of 1, 2, 3, and 4 wt pct Cu. Spec-trographic and chemical analyses are listed in Table I. After heat treatments and intermediate reductions, the alloys were cold rolled to 1/8 in. thick sheet. The final cold reductions amounted to 50, 40, 20, and 10 pct for the 1, 2, 3, and 4 pct Cu alloys, respectively. The specimens for creep testing were machined into tensile flats 7 in. long, 3/4 in. wide (at the shoulders), and 1/8 in. thick. The reduced sections were 1/4 in. wide and 21/2 in. long. After heat treatments at 540°C to an ASTM grain size of 1 1, the specimens were quenched into water at room temperature, then electropolished in a glacial acetic acid and perchloric acid solution. The gage marks were Knoop impressions placed 2 in. apart on the reduced sections. Blanks of 1 x 1 in, were cut from the same sheets for the age hardening runs and were given the same grain size heat treatments. The specimens and blanks were stored in a refrigerator at —40°C until needed for testing. Creep Measurements—Creep tests were conducted in a constant temperature room which was maintained at 26° ±1°C. A split-type, hinged, vertical-tube furnace was mounted on a horizontal track so that the specimen could be enclosed suddenly. Thus, it took about 15 min for the specimen to come within 5°C of the test temperature. The temperature variation along the specimen was controlled to about

Jan 1, 1958

By W. P. Sykes

In connection with a study of tungsten steels, Honda and Murakamil reported an investigation of the system iron-tungsten. This report included a tentative equilibrium diagram, photomicrographs of various alloys in the series, and qualitative measurements of hardness and magnetic properties. The present investigation originally had as its object the study of the hardness and tensile properties of a series of carbon-free iron-tungsten alloys. But, as the work progressed it became evident that the system possessed other features of interest, so the scope of the investigation was enlarged. Materials The iron powder used in preparing the alloys was obtained as follows : Iron oxalate was precipitated from a ferrous-sulfate solution by the addition of a solution of oxalic acid. This oxalate, after thorough washing, was ignited to iron oxide, which after reduction in hydrogen at about 1000" C. yielded iron powder containing but about 0.2 per cent. iron oxide and about 0.005 per cent. carbon. Tungsten metal in the form of powder, such as is used in the manufacture of filaments for incandescent lamps, was the other constituent. This material analyzed 99.8 per cent. tungsten and contained no carbon. The metal powders, in the desired proportions, were mixed by tumbling and formed, under pressure of 20 tons per sq. in., into rods 3/8 by 3/8 by 10 in. To form the alloys without contamination, sections of the pressed rods were placed on alundum slabs and heated in an alundum tube wound with a tungsten resistor and packed in heat-insulating material. While the resistor was heated, a stream of hydrogen was continually passed through both the tube and the case containing it. The tungsten winding on the tube, as well as the material heated within the tube, was thus protected from oxidation. Such a furnace will operate constantly for several days at temperatures up to 1500' C. and for shorter periods

Jan 1, 1926

The Palabora ore consists of two types, a carbonatite containing a low titanium magnetite, and a foskorite containing high titanium magnetite. The latter type also contains substantial quantities of phosphate. These two ore types are separately crushed and are kept separate through copper flotation and magnetic separation, for recovery of magnetite, with the low Ti magnetite being shipped as iron ore. High phosphate tailings are pumped to Foskor, a government-owned phosphate plant nearby. The Palabora plant was originally designed in 1966 for 33, 000 tpd using two 54 in. by 74 in. Allis -Chalmers gyratories; two 7 ft (2. 13 m) standards; four 7 ft shortheads; and six grinding lines, each using one 1,200 hp rod mill and two 1,200 hp ball mills. A fifth short- head and one more grinding line were added later; and with flowsheet improvements, the capacity of the plant was increased to 59, 500 stpd. A further plant expansion involving a third 54 in. primary and two 32 ft by 15 ft (9.75 by 4.6 m) 9,000 hp autogenous mills and another flotation section will bring its capacity to 87,500 stpd in 1977. These two fully autogenous mills are constructed entirely out of doors near the primary crusher, and the ground ore will be pumped a mile to the concentrator by a five-stage pumping system. The Palabora ore is mined with 6, 10, and 12 7d (4.6, 7.6, and 9. 2 cu m) shovels and hauled to the crushers in 110 and 165 st trucks. Sixty-five ton trucks were originally used, but the crusher dump pockets were designed for larger units. Dump pockets, crushers, feeders, and take- away belts are offset in order to keep the two ore types separate. Centerlines of the crushers are only 43 ft (13. 1 m) apart; but, because of their being offset, maintenance room is ample. Initially, more carbonatite ore than foskorite was mined; but as the pit has deepened, this ratio has been reversed. The Palabora ore breaks in the pit with a large amount of oversize boulders; and, with the use of larger trucks, a splitter has been installed on one side of each dump pocket to protect the crusher spider. Smaller trucks dump on the opposite side. To accomodate the splitter, the pocket edge has been raised one meter and moved back one meter. Dust is collected through a special vent shaft between the crushers and the mantle relining area; but even so, two high-pressure water spray headers with nozzles are installed under the edge of the pocket at each dumping position. While only one truck dumps at a time, sprays from all four headers are used to suppress dust. Any leakage of water through the seal around the upper bowl of the crusher is caught by a circular trough under the seal and directed to a sump pump. At this point, there

Jan 1, 1978

By E. A. Anderson

THIS seems to be the year for superlatives in A.I.M.E. meetings. The programs of the various Divisions and Institute committees offered an abundance of interesting and valuable information in the form of technical papers, symposia. and lectures to large and interested audiences. The Institute of Metals Division may hold up its head with pride for its own program. With six technical sessions, three full-day symposia, and the annual I.M.D. dinner, four meeting days were not enough and many simultaneous sessions were required. The quantity and quality of the paper were equaled by the numbers and interest of those attending.

Jan 1, 1944

By M. C. Manderson

Abstract-The sharp drop in world phosphate demand that took place in 1975, due to temporarily high prices, now seems to be reversing itself. And prices for both phosphate rock and phosphate fertilizers, which declined dramatically from their 1975 highs, now appear to be bottoming out. This recent price weakness has delayed new mining ventures that had previously been planned for completion by around 1980. While there is ample capacity to supply near-term rock needs, the stage seems to be emerging for a better supply-demand and price environment in the late 1970's and early 1980's.

Jan 1, 1978

By H. I. Aaronson, W. L. Winterbottom, G. M. Pound

On the basis of a statistical thermodynamic treatment of the data of smithz2 on the activity of carbon in austenite, a Darken and smith23 deduced that the interaction energy, wy, between carbon atoms occupying nearest-neighboring interstitial sites is repulsive. Aaronson, Domian, and pound 24 (ADP) have applied two different statistical thermodynamic analyses to the same data. One analysis, a composite of the treatments of Darken and Smith,23 Speiser and Spret-nak," and Kaufman, Radcliffe, and Cohen,26 primarily takes account of the positional entropic effects of wy in dilute solutions. The other, due to Lacher 27 and Fowler and Guggenheim," is also applicable to more concentrated solid solutions. The equations which both analyses produce for a,, can be rearranged so that, when In a,, is plotted against the appropriate function of the carbon content of the austenite, the plot will have a slope of unity when the correct value of wy is employed. Since no other constant need be simultaneously evaluated in order to make this determination, ascertaining the most probable value of wy from Smith's data on the bases of these analyses was an elementary problem in least-squares statistics. ADP demonstrated that the wy's obtained from both analyses vary with temperature in a statistically significant manner. McLellan, Gerrard, Horowitz, and sprague 29 (MGHS), on the other hand, have recently made a new statistical thermodynamic study of a,, and have concluded from this study that wy is independent of temperature. The equation which they derived for a,, is: where 0 = ratio of the mole fraction of carbon atoms to that of iron atoms, z = number of sites from which occupancy is excluded by the presence of a carbon atom and the repulsive character of its interaction with nearest-neighboring carbon atoms (z = zb in Ref. 24), and x = a geometric parameter taking account of overlapping of the exclusion shells of nearby carbon atoms. z is 1 when wy = 0 and 13 when wy = Q. For a given value of z there is a definite value of X . In the absence of an equation connecting z and X, MGHS computed geometrically the value of x at several integral values of z; x increases from 0.47 to 8.67 as z is increased from 2 to 5. MGHS fitted Eq. [I] to the a,, data of Ellis, Davidson, and Bodsworth30 at 925° and 1050°C and to the data of smithz2 at 1000°C. Confining z to integral values in the range 4 to 6, and also somewhat restricting AG, the values of z, X, and AG were sought which would minimize the average root mean square percent deviation, D, of the calculated a,, values from those obtained experimentally. Finding that D was almost always smallest when z = 5, they concluded that z is independent of temperature. Since:23, 26 where R = gas constant and T = absolute temperature, and w,, was considered to be similarly invariant, this statement cannot be strictly true. The temperature range employed, however, was small and the interval between the values of z tested was sufficiently large so that the results could still be consistent with a constant The statistical treatment of the a,, data applied by MGHS is clearly quite approximate. Since more than three values of a,, are available at each temperature, z, and X, and AG have actually been over deter mined, rather than underdetermined as their treatment implies. In order to make more complete and exact use of the a,, data, we shall employ the Legendre Method of Least Squares, as described by Whittaker and Robinson. This technique permits conversion of the overdetermined system of equations produced by the substitution into Eq. [I] of the available pairs of a,, and 0 data at a given temperature into an even determined system. Solution of the latter system yields the most probable value of each of the three constants. This procedure will allow a statistically rigorous test to be made of the conclusion drawn by MGHS in respect of the temperature dependence of In order to obtain linear iiequations of conditions" from Eq. [I], the exact values of the three unknown constants are replaced by G/RT = =/RT + y, where z, i, and aG are estimated values of the constants. This allows the conditional equations to be written in the form:'l where fi -f6, i, =/RT) = the right-hand side of Eq. [I] for a given pair of ayi and values. When applied to Eq. [I], these relationships become: The i equations of condition are reduced to three "normal equations" in the manner discussed by Whittaker and Robinson. The normal equations are then solved for a, ß, and y. Both of these operations are most conveniently performed by means of the computer technique of Winterbottom and Gjostein.32 This technique was used iteratively until the value of 0 (the most sensitive of the three correction factors) became 10.01. (Until they became small, a, 8, and y

Jan 1, 1969

By W. N. Tanner

THE wire-rope hoisting conditions at the mines of the Anaconda Copper Mining Co. in Butte, Mont., are very severe because of the conditions under which it is necessary to operate. A study was made, in the fall of 1919, for the purpose of changing these conditions in order to obtain longer rope and sheave life. There are at present in Butte, hoisting engines equipped with wire rope as follows: 9 main hoists using flat ropes. 7 main hoists using 1 ½ in. rope. 5 main hoists using 1 ¼ in. rope. 6 shafts using 1 1/8 in. rope. 2 chippy hoists using 1 ¼ in. rope. 9 chippy hoists using 1 1/8 in. rope. 15 shafts using 1 in. rope. 7 shafts using 7/8 in. rope. 1 shaft using ¾ in. rope. In addition, there are a number of sinking engines using 5/8 in., ¾ in., 7/8 in., and 1 in. (15.9, 19, 22, and 25 mm. rope.) The flat ropes run from ½ by 7 ½ in. (12.7 by 190.5 mm.) to ½ by 4 ½ in. (9.5 by 114 mm.) and are used on the older hoisting-engines, which are set comparatively close to the shaft. Because of swelling and moving ground at some of the shafts, these ropes present certain difficulties in operation that are more pronounced than with round ropes; they are, therefore, being superseded by round ropes as newer engines are being installed. They are, however, much more easily inspected and repaired than round ropes.

Jan 2, 1922

By AIME AIME

AN IMPORTANT innovation in Alabama in the semi-longwall type of coal mining as applied to low-dipping thin seams has been introduced by the Galloway Coal Co., mining the Mary Lee high-ash seam which averages only 26 in. thick. The property is their Mill Creek mine, near Carbon Hill, Ala., in the Warrior coal field. The results, although not final, indicate that roof conditions can be successfully handled by this method without loss of walls, with a rate of recovery much above that possible with the room-and-pillar method, and at the low production cost necessary for this grade of coal. Double cross-headings are turned off a main haulageway on 60-ft. centers and driven 40 ft. wide with thin chain pillars between. These double entries are turned at 525-ft. intervals, leaving a solid block of coal 440 ft. across; that is, from the outside of one pair of headings to the inside of the next pair is 440 ft. The bottom, which consists of hard fireclay, is taken in these entries 8 ft. wide and 4 ft. deep for head room.

Jan 1, 1934

By Bunting S. Crocker

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

Jan 1, 1953

By Bunting S. Crocker

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

Jan 1, 1953

By Robert R. Walden, LeMont West

K ENTUCKY'S first acid-grade fluorspar flotation Kmill, shown in Fig. 1, was placed in operation Aug. 1, 1952, by the Pennsylvania Salt Manufacturing Co. at Mexico, Ky. During 1951 a critical shortage of acid-grade fluorspar developed and it became necessary for the company to produce part of its fluorspar requirements. In 1951 the company purchased the Crider and Lennen metallurgical fluorspar mill and an adjacent stockpile of tailing at Mexico, Ky. The tailing, assaying about 26 pct calcium fluoride, originated from log washer and jig mill concentrators in the area. As the mill did not have a fluorspar dryer or convenient rail facilities, the company leased the Burleson concentrator at Marion, Ky., where these facilities were available. At Marion, Ky., on the Illinois Central Railroad, dry acid-grade fluorspar is transported direct to the company's Calvert City, Ky., hydrofluoric acid plant. The Mexico flotation plant was reconditioned and designed to produce acid-grade fluorspar from the jig tailing stockpile. The new plant flowsheet is shown in Fig. 2. The ore feed, old tailing, which contains 10 pct of waste rock and coarse gravel assaying approximately 8 pct CaF,, is bound together by sticky clay

Jan 1, 1955

By A. E. Seaton

In this paper it will be my endeavor to trace the development of the marine engine and its appurtenances, and the general progress that has taken place in marine engineering generally during the past fifteen years, rather with the object of attributing much of it to the successful efforts of the makers of iron and steel, thereby rendering my thanks to them for the valuable service they have contributed to my profession, than to recount what has been achieved by the skill, ability and perseverance of my colleagues in shipbuilding and engineering. The engineer has been at all times dependent on the ironmaster, and on more than one occasion his progress has been barred by the inability of the latter to overcome some obstacle to success in his manufacture. But on the other hand, it is equally true that the ironmaster has often been ahead of the engineer, and has been prepared, long beforehand, to supply that which would make a success for the marine engineer if he would avail himself of it.

Jan 1, 1891