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By Michael A. Zappia

INTRODUCTION The San Manuel mine of Magmacopper Company is located approximately 72 km northeast of Tucson, Arizona. The mine, concentrator, smelter, electrolytic refinery and continuous rod casting are located on the west side of the San Pedro River valley. Ore Production: 58 1W mtpd Production To Date: Over 340 000 000 mt since 1956. Ore Grade: 0.70% net sulfide cu. As conceived at the present time, the San Manuel ore body is the lower half of what was originally an elliptical (In cross sect ion) cylinder same 2400 m long with major and minor cross sectional axes of 1500 m and 760 m. After emplacement, the cylinder was diagonally sliced along its long dimension by the San Manuel fault, which separated it into halves, moving the upper half same 2500 m down dip and bringing a wedge-shaped blanket of tertiary conglomerate into position over the lower segment of the cylinder. The lower segment is the San Manuel ore body. The upper part, moved down dip, is the Kalamazoo ore body. Originally, these were one ore body. The San Manuel ore body is a structurally weak and highly fractured quartz monzonite, the principal ore minerals being chalcopyrite, with small amounts of molybdenite. The present mining method is full gravity diagonal retreat panel caving by block. This is shown on Figure 2 , (Block Sequence 56, 57, 58, 59 --on the following page.) The ore mined is hoisted to the surface for transportation to the concentrator. BLOCK CAVING MINING SEQUENCE HISTORY The San Manuel mine was planned and developed as a full gravity block caving mine. It continues as a full gravity block caving mine. Since its initial start, the major changes in mining have been in the areas of ground support, rate of draw and undercut sequence. Ground weight resulting from block caving plays a dual role. It can cause extensive damage to ex- traction openings, yet it is necessary to break the ore into easily handled size. The ground- weight problem resolves itself into finding ways to make it manageable. The initial early ground support in the production areas consisted of timber sets and various combinations of steel and timber sets. This sup- port system proved unsatisfactory and a new support system utilizing concrete was developed. Concrete support, 30 cm to 45 cm thick, proved satisfactory. Rates of draw were experimented with. A draw rate of 45 cm to 61 cm (32 to 45 mt) per draw point per day was determined as optimum in con- trolling ground weight. Undercut sequence or mining sequence was recognized as a major contributing factor to ground weight. This is the most difficult factor to change. The lead time from primary development to undercut can be as much as two years. Initial Checkerboard Pattern The initial undercut sequence was in a checker- board pattern (Figure 2). Blocks Sequence 1, 2, 3, 4, 5, 6--next page.

Jan 1, 1981

By Lawrence Adler

For openings in bedded rocks, analyzed by simple beam theory, it has been shown that roof loads can be shifted from one support to another.' This transfer is effected by controlling the relative deflection of adjacent supports. If this action is used to reduce loads in the props or bolts and increase them in the ribs or pillars, it could be productive of economies averaging 30 pct. Such savings can be realized by: 1) using smaller artifical supports such as props and bolts, and 2) increasing loads in the rib to induce eventual failure for longwall caving. Since the rib is relatively load-insensitive, it provides a natural load sink. However, caution must be used in applying such permissive deflections, for while the supports are being handled properly, the stresses in the roof itself are very sensitive to such manipulations.'

Jan 5, 1960

By Thomas H. Clagett

ThE Pocahontas coal field comprises the area in Tazewell County, Va. and Mcrcer and McDowell counties, W. Va., in which Nos. 3 and 4, Pocahontas seams of bituminous coal are mined. It is a mountainous region with ridges rising 300 to 1000 ft. (91 to 305 111.) above the very narrow valleys and, except for some poorly kept farms along the tops of the ridges and the area occupied by the mining and coking plants and villages, is wooded throughout. First shipments were made from Pocahontas in 1883. There are now about 100 mines, 95 per cent. of which are drift mines. At the end of 1920, approximately 298,000,000 gross tons of coal had been produced and 35,000 acres of coal mined. No. 3 seam is from 4 to 10 ft. (1.2 to 3 m.) thick, but the change is gradual and the thickness fairly uniform in any particular locality. It has a streak of bone, about 2 in. (5 cm.) thick, to which the coal adheres on both sides. In parts of the field this bone is replaced by hard slate, usually from 2 to 4 in. thick. The seam has a slate or fireclay bottom and a draw-slate top. No. 4 seam is from 65 to 80 ft. (19.8 to 24 m.) above No. 3 seam, and of minable section only in the western part of the field. It is from 3 to 8 ft. thick and has two streaks of bone, cach from 1 to 2 in. thick. It has a fireclay or slate bottom and is overlaid by from 4 to 24 in. of a laminated coal and slate, locally termed "black rash," which is separated from the coal seam by about 3 in. of slate. This "black rash" averages about 2.5 per cent. ash, but contains some clean coal. The coal in both seams is quite soft. Cleavage sufficieiltly well defined to exert an influence on the mining is rare. Overlying strata arc slates, shales, and sandstones, the sandstones predominating, extending from 30 to 1000 ft. above No. 3 seam. The strata directly above the draw slate of No. 3 seam, or "black rash" of No. 4 seam, and up to the first sandstone are usually slates or shales, about 10 ft. thick over the.

Jan 1, 1923

By Thomas H. Clagett

ThE Pocahontas coal field comprises the area in Tazewell County, Va. and Mcrcer and McDowell counties, W. Va., in which Nos. 3 and 4, Pocahontas seams of bituminous coal are mined. It is a mountainous region with ridges rising 300 to 1000 ft. (91 to 305 111.) above the very narrow valleys and, except for some poorly kept farms along the tops of the ridges and the area occupied by the mining and coking plants and villages, is wooded throughout. First shipments were made from Pocahontas in 1883. There are now about 100 mines, 95 per cent. of which are drift mines. At the end of 1920, approximately 298,000,000 gross tons of coal had been produced and 35,000 acres of coal mined. No. 3 seam is from 4 to 10 ft. (1.2 to 3 m.) thick, but the change is gradual and the thickness fairly uniform in any particular locality. It has a streak of bone, about 2 in. (5 cm.) thick, to which the coal adheres on both sides. In parts of the field this bone is replaced by hard slate, usually from 2 to 4 in. thick. The seam has a slate or fireclay bottom and a draw-slate top. No. 4 seam is from 65 to 80 ft. (19.8 to 24 m.) above No. 3 seam, and of minable section only in the western part of the field. It is from 3 to 8 ft. thick and has two streaks of bone, cach from 1 to 2 in. thick. It has a fireclay or slate bottom and is overlaid by from 4 to 24 in. of a laminated coal and slate, locally termed "black rash," which is separated from the coal seam by about 3 in. of slate. This "black rash" averages about 2.5 per cent. ash, but contains some clean coal. The coal in both seams is quite soft. Cleavage sufficieiltly well defined to exert an influence on the mining is rare. Overlying strata arc slates, shales, and sandstones, the sandstones predominating, extending from 30 to 1000 ft. above No. 3 seam. The strata directly above the draw slate of No. 3 seam, or "black rash" of No. 4 seam, and up to the first sandstone are usually slates or shales, about 10 ft. thick over the.

Jan 1, 1923

By G. R. Belton, Y. K. Rao

A galvanic cell, using liquid MgCl2 or MgC12-CaC12 mixtures as the electrolyte, has been used to determine activities in Mg-A1 liquid alloys between 700' and 880°C. The incovporation of a chlorine electrode in the cell also allowed measurements to be made of the standard free energy of formation of MgCl2(l). The results are shown to be in good agreement with thermo-chetrlical values from the literature, and this is taken as evidence that the small, known solubility of magnesium in iMgCl2 introduces no significant error in galcanic cell measurements. Within experimental error, the activity coefficients and relative partial molar enthalpies at 800°C are shown to be represented by the following "subregular" solution equations: logy~~ =-0.68(1 -xMgj3 log yAL =-1.02(1 - XMf + 0.68(1 - XMf H.M = -4400(1 - %)3 / cal Mg 7Ag' HZ =-6600(1 - xA1)' + 4400(1 - XMf cal SCHNEIDER and toll' have used a transpiration technique to measure the vapor pressures of magnesium over Mg-A1 alloys between 544" and 594°C. amsstad,' however, has since suggested that the extrapolation to zero flow rate, used by these authors in interpreting their apparent pressure vs flow rate data, gives unreliable results. Rogers, Tomlinson, and Richardson,3 in interpreting the results of solution equilibria between Mg-A1 alloys and liquid MgC12, also considered the measurements of Schneider and Stoll to be unreliable and preferred to derive activities for the alloys from the earlier boiling point determinations of ~eit~ebel~ and the partial molar heats recommended by Kubaschewski and ~atterall.~ These latter heats were based substantially on the early calorimetric work of Kawa-kami but, unfortunately, his work on other systems has sometimes been found to be inaccurate.7 Rogers et al., in the above-mentioned paper,3 tentatively concluded that the most likely species responsible for the limited solubility (0.3 mole pct at 800°C) of magnesium in MgC12 were Mg° (neutral) and Mg2++. Two more recent studies8,9 have supported Mg2++ as the soluble species. In the present study, activities in Mg-A1 alloys have been determined by means of a galvanic cell involving liquid MgCl2 or MgC12CaCl, mixtures as the electrolyte. Since the reactive nature of magnesium precluded simple Faraday yield experiments, a chlorine electrode was incorporated in the cell in order that the performance of the cell could be checked by measurements of the heat and free energy of formation of MgC12. This procedure was considered necessary since it has been suggested1' that the solubility of a metal in a molten salt might introduce electronic conductivity; also, previous determinations of the standard electrode potential for MgC12 differed by as much as 70 mv11-13 EXPERIMENTAL Materials. Analyses of the materials used in preparing the alloys and the electrolyte mixtures are presented in Table I. The alloys were prepared by induction melting weighed amounts of the metals in a graphite crucible held under an argon atmosphere. Pure anhydrous magnesium chloride was prepared by heating the mixture MgCl, . 6H20 +NH4C1(1:1) to 650°C, followed by melting under dry argon. The melting point of the dry MgC12 was found by differential thermal analysis to be 714.8oC, which compares well with the accepted value of 714C.14 This agreement was taken to be an indication of the high purity of the dried salt. Table I. Compositions of Materials, wt pn Impurity Mg Al MgCI2 CaCI, Ba - - 0.005 Ca - - 0.010 Cu 0.02 0.02 Fe - 0.10 0.001 0.010 Pb 0.01 - 0.001 0.005 Mn 0.15 0.001 Si - 0.10 Sr - - 0.005 MBSO* - 0.040 ARGON TUNGSTEN CHLORINE LEAD—411 11 II ALUMINA 'A I / ave GRAPHITE on irA~_ ( ^"l ROD MAGNESIUM S>ILICA~^ /, OR ALLOY-. I /A] ___________ -«^- _ ELECTROLYTE I y FRITTED DISCS Fig. l—Arrangement of chlorine and metal electrodes in electrolytic cell.

Jan 1, 1970

By W. C. Maurer

A drilling-rate formula for roller-cone bits is derived from rock crater-ing mechanisms. This formula holds for "perfect cleaning", which is defined as the condition where all of the rock debris is removed between tooth impacts. Under these conditions, the drilling rate is directly proportional to the rotary speed and to the bit weight squared, and inversely proportional to the bit diameter squared and to the rock strength squared. Under imperfect cleaning conditions, such m those usually present in field drilling, re-grinding of the cuttings occurs under the bit, and the drilling rates fall below those for perfect cleaning. INTRODUCTION Drilling with a roller-cone bit consists of two fundamental operations: the formation of craters under the bit teeth, and the removal of the broken rock from the craters. To delineate the mechanisms of drilling, it is first necessary to understand the mechanisms involved in the formation of individual craters and then to relate these mechanisms to the over-all drilling operation. CRATER FORMATION When a bit tooth impacts a rock, the rock is elastically deformed until the crushing strength of the rock is exceeded, at which time a wedge of crushed rock is formed below the tooth1, 2 as shown in Fig. 1. As additional force is applied to the tooth, the crushed material is compressed and exerts high lateral forces on the solid material surrounding the crushed wedge.- en these forces become sufficiently high, fractures are initiated below the tooth and propagate to the free surface of the rock. The trajectories of these fractures intersect the principal stresses at a constant angle,'s4, as predicted by both Mohr's and Griffith's theories of failure. It has been experimentally 3,1 .6 that the volume of an impact crater, V,, varies as where Ec is the total energy imparted to the rock during the formation of the crater, and E is the threshold energy required before cratering is initiated (Fig. 2). Indexing experiments' have shown that, when there is a second free face to which the crater can fracture, a larger volume of material is removed and that above a transition zone there is a linear relationship between V, and Eo., which has a slope equal to the relationship for the single crater. The beneficial effect of indexing tends to offset the threshold energy requirement to a large extent, as shown in Fig. 2, and the relationship is closely represented by Rate-of-loading effects have been used to explain the nonproportional relationships usually obtained between drilling rate and rotary speed. The data4, in Fig. 3 show that the Ec/Vc parameter remained constant for craters produced with impact velocities ranging from 10 to 8,000 ft/sec. Since no rate-of-loading effects were detected over this wide range of loading rates, it seems improbable that there would be detectable rate-of-loading effects present in the small changes in impact velocities produced by varying the rotary speed. The writer has shown (Fig. 4) that, for craters produced in rock by impacting spheres, where X is the depth of penetration. Simon2 has shown that this relationship also holds for craters produced in rock by penetrating wedge-shaped chisels. The energy used in crater formation is

By J. G. Baragwanath

CUBA'S lateritic iron ores, occurring on the northern coast of that island, though known to the Spaniards. did not receive any general attention until the close of the Spanish-American War. Preliminary studies of these deposits at that time indicated that the limonite contained small percentages of nickel and chrome, and it ryas subsequently discovered that although the average nickel content was only approximately 0.6 per certain limited areas of ore were suffiniently high in nickel to be considered commercial provided a suitable process fur its recovery could be developed. Early in 1937, Pardners Mines Corp. began a series of investigations which led to the development of such a process. Coincidentally, this company acquired option' on certain properties in the Levisa Bay and Moa Bay regions and conducted a systematic campaign of exploration by test pits and hand-auger holes in the course of which an aggregate of 90.000.000 tons of ore was developed in eight deposits.

Jan 1, 1945

By AIME AIME

ON the headwaters of the Hudson Riser, in a sparsely populated area of the north woods at Tahawus, N. Y., thirty miles from the nearest railroad, is the Maclntyre property of National Lead Co. Operations involve the mining and processing of titaniferous iron ore to produce ilmenite concentrate, and magnetite concentrate containing vanadium, for use in the production of iron, titanium, and vanadium in the war program. Accordingly, Maclntyre at present is classed as a war plant. More than 3000 long tons of ore is ruined daily by opencut methods and treated in an ore concentrating plant.

Jan 1, 1943

By J. E. Dinger, Ross Gunn

Faraday's Law of magnetic induction states that the e.m.f. acting around a closed circuit is equal to the time rate of change of flux through that circuit. If the closed circuit is a conducting medium, there will flow an electric current which is propor¬tional to the conductivity of the material and the time rate of change of the flux. Electrical currents so induced in a con¬ducting material are often referred to as eddy currents. Eddy currents are thus induced into a conducting material by (1) a time rate of change of the magnetic field through the material, by (2) a relative motion of the conducting material and a magnetic field, or by (3) a combination of (1) and (2). Because the magnitude of the induced eddy currents is dependent upon the resistance of the equivalent conducting path, these currents offer a method of detecting electrical discontinuities in a

Jan 1, 1945

Joseph William Richards, was born in Oldbury, Worcestershire, England, on July 28, 1864, of English-Scotch parents, Joseph and Bridget (Harvey) Richards. In 1871, he came to this country with his parents. He attended the public schools of Philadelphia, Pa., where his father was a manufacturing metallurgist, and in 1882, graduated from the Central High School with the degree of A. B. The same year he entered Lehigh University, from which he received the degree of A. C. (Analytical Chemist) in 1886; M. S. in 1891, and Ph. D. in 1893. His connection with the university was continuous from 1887 until his death, and he therefore adds another to the list of names of men Prominent in the A. I. 11. E. who have helped to develop Lehigh University to the important position which it now occupies in mining and metallurgical education in respect of world-wide fame. It is fitting to recall in this connection that, Eckley B. Coxe, a fonnder of the A. I. M. E., four times its vice-president,, and president in 1878-79, was literally a pillar of the university for several years prior to his death; Thomas M. Drown, founder, vice-presi(lent, secretary for 11 years, honorary member, and president in 1897, was president of Lehigh University from 1895 to 1904; and Henry S. Drinker, founder, manager, vice-president, and honorary member, was president of the university from 1905 to 1920. Doctor Richards was no less active in both institutions. With the exception of the years 1886 to 1887, when he was superintendent of the Delaware Metal Refinery of Philadelphia, his professional life was devoted to the upbuilding of the metallurgical department of the university, in part through his continuous services as a member of the teaching staff, and in part through the international renown won for the department by his achievevements as author, investigator, expert and rerognizcd authority in metallurgy, especially the metallurgy of aluminum, metallurgical-chemical calculations, and electrometallurgy. He was assistant instructor in metallurgy and blowpiping in 1887; instructor, 1888-90; instructor in metallurgy, mineralogy and blowpiping, 1890-97; assistant professor, 1897--1903; professor of metallurgy, 1903-1921. He died at his home in Bethlehem, Pa., on Oct. 12, 1921, of heart failure. To the A. I. M. E., for a continuous period of a dozen years, his services were equally loyal and effective. He was vice-president, 1910-11, and 1916-17; councilor, 1912; dirctor, 1913-15; chairman of the Iron and Steel Clommittee, 191-1-21; he was one of the framers of the Insti-

Jan 1, 1922

Joseph William Richards, was born in Oldbury, Worcestershire, England, on July 28, 1864, of English-Scotch parents, Joseph and Bridget (Harvey) Richards. In 1871, he came to this country with his parents. He attended the public schools of Philadelphia, Pa., where his father was a manufacturing metallurgist, and in 1882, graduated from the Central High School with the degree of A. B. The same year he entered Lehigh University, from which he received the degree of A. C. (Analytical Chemist) in 1886; M. S. in 1891, and Ph. D. in 1893. His connection with the university was continuous from 1887 until his death, and he therefore adds another to the list of names of men Prominent in the A. I. 11. E. who have helped to develop Lehigh University to the important position which it now occupies in mining and metallurgical education in respect of world-wide fame. It is fitting to recall in this connection that, Eckley B. Coxe, a fonnder of the A. I. M. E., four times its vice-president,, and president in 1878-79, was literally a pillar of the university for several years prior to his death; Thomas M. Drown, founder, vice-presi(lent, secretary for 11 years, honorary member, and president in 1897, was president of Lehigh University from 1895 to 1904; and Henry S. Drinker, founder, manager, vice-president, and honorary member, was president of the university from 1905 to 1920. Doctor Richards was no less active in both institutions. With the exception of the years 1886 to 1887, when he was superintendent of the Delaware Metal Refinery of Philadelphia, his professional life was devoted to the upbuilding of the metallurgical department of the university, in part through his continuous services as a member of the teaching staff, and in part through the international renown won for the department by his achievevements as author, investigator, expert and rerognizcd authority in metallurgy, especially the metallurgy of aluminum, metallurgical-chemical calculations, and electrometallurgy. He was assistant instructor in metallurgy and blowpiping in 1887; instructor, 1888-90; instructor in metallurgy, mineralogy and blowpiping, 1890-97; assistant professor, 1897--1903; professor of metallurgy, 1903-1921. He died at his home in Bethlehem, Pa., on Oct. 12, 1921, of heart failure. To the A. I. M. E., for a continuous period of a dozen years, his services were equally loyal and effective. He was vice-president, 1910-11, and 1916-17; councilor, 1912; dirctor, 1913-15; chairman of the Iron and Steel Clommittee, 191-1-21; he was one of the framers of the Insti-

Jan 1, 1922

By Harlowe Hardinge, R. C. Ferguson

Laboratory and plant data covering 12 different operations show that lower than "standard" ball mill speeds increase grinding efficiency. In the case of high pulp-level mills, the gain is so great that the increase in capital cost of the larger lower speed mill will pay for itself in less than a year's time.

Jan 1, 1950

By H. W. Morse

THE CHAIRMAN (H. W. MORSE).-Gentlemen, for the first time in the history of the American Institute of Mining Engineers, we have a full session on the subject of leaching-especially on the leaching of copper ores.. This branch of metallurgy is rather in its infancy. We have had one big plant running for a year or more, and pretty successfully; and another one which will be started in February or March, the New Cornelia. We have another one at Garfield for the Utah company, which is just designed and construction begun; and from several other points come reports of interesting test work. The first paper on the program this evening is one of some of the test work on the New Cornelia ore by myself and Mr. H. A. Tobelmann. LAWRENCE ADDICKS, New York, N. Y. (written discussion).-The Ajo proposition has had, as I understand it, two great advantages: it has not been obliged to measure leaching against flotation or some other metallurgical process, and it deals with a copper content high enough

Jan 12, 1916

By D. C. DERINGER

METALLURGISTS and mill operators appreciate, in a general way, .the economic or commercial relationship between recovery and grade of product but few have correlated in detail fluctuating metal prices with the metallurgy. They know that many factors such as metal price, dedusiions for returning charge, transportation, percentage of metal in head ore, etc.; all affect the commercial relationship between recovery and grade of product. In practically every mill some one at some ' time has calculated the amount of recovery that can be sacrificed for each percentage increase in grade of concentrate and from this calculated ratio the commercial or ., economic tail has been established. Such a ratio finally becomes a well-entrenched rule by which the progress of the metallurgy is judged.. When once established, these-, ratios are but seldom corrected to changes of metal prices for each possible grade of concentrate that may be produced. The calculation is tedious when often repeated. Other methods have been suggested and used to correlate the commercial with the metallurgical phase. A factor .derived ,by multiplying the grade of concentrate by recovery is often used, but is of little importance because it does not possess a single commercial quantity. The percentage monetary recovery calculated by dividing the gross value of the ore into the net smelter return gives a good economic index and, is to be commended, but it is difficult to correlate exactly and quickly the work of each day on a commercial basis.

Jan 1, 1931

By J. A. Richardson

J. A. RICHARDSON[t]-Mr. Stuckey's account of the geology of the kaolins of North Carolina suggests that they exhibit some features similar to those of the Malay Peninsula. About one half of British Malaya (some 26,000 square miles) is occupied at the surface by igneous rocks, of which granite and its associated derivatives constitute fully 90 pct. The average granite is a biotic variety. It is predominantly potassic but carries, in addition, notable amounts of soda and lime so that albite and albiteoligoclase are generally common constituents of the rock. The minor derivatives from granite "magma" likewise include potassic and sodi-potassic members. Decomposition of the granitic rocks has led to the formation of kaolin in many localities, but, in general, the deposits are rather small and, at present, none is exploited commercially. True primary or residual kaolins (as defined by Stuckey; p. 48 Vol. 173, p. 2 TP 2219) have been developed from feldspar veins, pegmatites and from other granitic differentiates which cut through the larger masses of granite; they are usually small and narrow. Kaolin has been derived from aplite, quartz porphyry, quartz micro-porphyry, felsites and microgranites belonging to a complex known as the Chegar Perah-Pulai intrusion in northwest Pahang and southwest Kelantan. The bulk of the component rocks has been more or less strongly sheared. Kaolin-like end products have been formed from both sheared and nonsheared varieties. Kaolin has been formed also by the complete breakdown of trondjemite flanking the northwestern and western margins of the complex Bukit Ranjut syenite intrusion in northwest Pahang. These trondjemites range in composition from essentially quartz-free albiteplagioclasites to quartz-free microcline rocks; they are, in effect, the syenitic homologues of white sodic and potassic granite-aplites. The granites in Malaya have generally undergone more or less severe surficial weathering leading to the production of a spongy and rather cellular or a soft "mealy" material (sandy clay) down to depths of about 100 ft; in places the granite is strongly decomposed down to 150 ft. There has occurred, therefore, widespread partial kaolinization, but, in general, no extensive deposits of good quality kaolin seem to have been formed by these processes of decay. Frequently, the katamorphic processes of desilicification have advanced beyond the production of kaolin towards the formation of lateritic end products. The more important Malayan kaoiins seem to be confined to the western alluvial plains, particularly in areas either where granite crops out at surface, or where it forms the bedrock of the alluvium. There are numerous excellent examples around Gopeng and elsewhere in the Kinta Valley of Perak. Such kaolins belong to Stuckey's groups of secondary deposits, and, like them, these also may be bedded. Field observations have established the existence of what for want of a better name is referred to by the Malayan Geological Survey as "granite wash." It is believed that granite, undergoing extensive deeply penetrative and probably rather rapid destruction under humid tropical weathering (temperature around 90°F; rainfall about 100 in. per annum; high humidity) "flowed," was moved by "creep" and, in addition, probably slumped in large waterlogged masses of sandy clay from topographic salients onto the adjacent flats. Either during this period of transmigration or later when the sludged granite detritus, after having more or less come to rest, was attacked by flowing water, differential sorting and erosion took place with the result that the resorted debris was moved and later deposited as bedded alluvium. In this way, banded "granite wash" was produced comprising alternating bands of clayey sand (resorted granite sludge) and white

Jan 1, 1947

By F. P. Haver, M. M. Wong

The Federal Bureau of Mines has developed a leach= electrolysis method for recovering lead from galena (PbS). Flotation concentrate is leached with ferric chloride (FeCl3) solution to obtain lead chloride (PbC12) and elemental sulfur, PbS + 2FeC13 -+ PbC12 + 2FeC12 + Sº. Over 99% of the P~S in the condentrate is converted to PbC12. The leach solution, composed of brine (NaC1) and FeC13, dissolves the PbC12 as it is formed, leaving a residue of elemental sulfur plus gangue. Very pure PbC12 (>99.9%).is crystallized from the leach filtrate on cool- ing and is subsequently electrolyzed in a fused-salt bath to obtain molten lead metal (>99.9% Pb). Spent leach solution (FeC12) is regenerated for further use by reaction with the chlorine obtained in the electrolysis of lead chloride, and other values, such as Ag, Cu, Zn, and Sº, are recovered by additional treatments of the leach solution and leach residue. This paper discusses results and conditions of the leaching and electrolysis operations.

Jan 1, 1976

By KENNETH CITTISGHAM

During the year 1945, the total number of wells drilled in Ohio, including the. non¬productive wells, was 1034. For the 10-year period ending with 1945, the average completions per year were 1125, the least number being 910 in 1938, and the maximum 1561 in 1941. The proportion of dry holes in 1945 was 37 per cent?about the same as for the year 1944 but exceeding the 10-year average, which was 34 per cent. The number of oil wells completed exceeded the number drilled in 1944 by about 30 per cent, the average initial per well being the same in both years. Gas-well completions were approximately 10 per Cent less than in 1944, with the average initial per well about 5 per cent greater. The difference in number of oil and gas wells in the two years was in Berea and Clinton completions, as the number of oil wells increased about 30 per cent, owing to greater activity in the two sands named, while the number of gas wells decreased because of fewer completions therein.

Jan 1, 1946

By G. J. Heuer, J. N. Dew, G. C. Clark

The effect on well behavior of partial permeability barriers, changes in producing rates and well spacings have been calculated through use of a radial, unsteady-state, two-phase-flow mathematical model. This model allows for variations in permeabilities and porosities with distance from the wellbore. The numerical methods necessary to solve the problem require use of a highspeed digital computer, in this case an IBM 704. In each instance, pressure and saturation gradients, gas-oil ratios and recoveries around a well producing by solution-gas drive have been calculated as a function of time and distance from the wellbore. Comparisons are made to show the effect of changes in producing rate and of varying permeability near the wellbore and out in the formation on the production and pressure history of the well. The effect of different well spacings on producing rate and ultimate recovery is considered. Although the mathematical description of the reservoir is simplified in comparison to an actual reservoir, the results do give some insight into the difficult problem of spacing and proration in a heterogeneous solution-gas-drive reservoir. Results show that for the cases considered spacing has little effect on ultimate recovery and that permeability barriers removed from the well decrease producing rate for a period of time but have only a small effect on ultimate production. INTRODUCTION The effect of spacing and producing rate on the production characteristics and ultimate recovery of a well producing by the solution-gas-drive mechanism have been topics of interest to the oil industry for many years and the subject of many hearings before state regulatory bodies. The problem is very complicated because of the difficulty of simulating in the laboratory a reservoir producing by the solution-gas-drive mechanism under anything like normal field conditions. This paper gives the initial results of a study aimed at determining the effect of these factors on production from reservoirs simulated by a mathematical model. Two-phase, unsteady-state equations are used to calculate pressures, saturations, and rates of oil and gas flow as a function of time and distance from the wellbore. The model is radial, and reservoir properties may be varied with distance from the wellbore. A decrease in permeability at a given distance results in a low-permeability ring concentric with the wellbore through which all the fluids from more distant portions of the reservoir must flow. The current mathematical model allows for variation of permeability, porosity and saturations as a function of distance from the well. Any desired drainage radius for the well may be selected. Drainage radii of 745 and 1,053 ft, which correspond to 40- and 80-acre spacing, were chosen for the calculations which follow. A large amount of the input data has been held constant. No investigation was made into the effect of changes in fluid properties and relative permeability characteristics on oil and gas production behavior. Fluid properties and relative permeabilities are essentially those of the White Mesa portion of the Greater Aneth area in southeastern Utah. A recently concluded hearing before the Utah Oil and Gas Conservation Commission to determine spacing for the field aroused interest in the possible effect of permeability constrictions located some distance from the well on production and recovery. This interest stimulated the present study. In this paper, no attempt has been made to simulate any field production characteristics exactly. Only the solution-gas-drive mechanism has been investigated, and production by liquid expansion has not been considered. This investigation is exploratory in nature but does give some insight into the problems of spacing and proration in a solution-gas-drive field having permeability variations. THE MATHEMATICAL MODEL The mathematical model is similar to that proposed by West, Garvin and Sheldon.' Approximately the same techniques have been used in arriving at a solution. The equations are nonlinear, partial differential equations which have been known for some time. The particular equations used for this study neglect the effects of capillary pressure and gravitational forces. The advent of modern digital computing equipment has made their solution practical. Basically, the method solves for pressure, saturations, and oil and gas flow rates as a function of distance and time. In our model, distance to the

By C. S. Allen

With increased coal production and mechanization of coal mines many transportation problems arise. The main objective is to bring the coal to the tipple or dump it as quickly as possible. Larger and faster haulage locomotives or two locomotives connected in tandem have been adopted to solve the problem. Motor equipment, gears and control systems of locomotives to be placed on these hauls are carefully studied to meet these requirements, by an analysis of motor characteristic curves with the corresponding tractive effort curves, in order to determine the correct load for maximum efficiency and to determine the approximate schedule. Control equipment is usually considered to be the controller and resistance. Sometimes accelerating curves are calculated in order to obtain smooth starting in the least possible time without slipping the wheels. Train control is very important, but why limit this to acceleration ? Deceleratien-or-retardation of the locomotive is a part of the train control and should be given as much consideration as the control for acceleration. Types or Braking Common braking systems for mine locomotives utilize screw-type mechanical brakes, dynamic braking and the air brake. For gathering and haulage locomotives, the srrew-type mechanical brake is regarded as standard. A brake of this kind cannot be applied without much effort and lapse of time. Motormen, like other people, follow the path of least resistance, and therefore have their brakes ready for instant application—which means that the shoes are virtually dragging. Even then, a motorman bucks and plugs the motors to gain time and avoid as much physical effort as possible. "Bucking of motors" consists in turning the reverse cylinder of the controller to parallel position corresponding to the opposite direction of travel while the locomotive is in motion, without applying power from an external source. When this occurs, both motors tend to act as series generators, because the armatures are revolving and there is residual magnetism in the field circuit. Owing to slight differences in motor manufacture, one of these will generate a slightly higher potential and will overcome the generated voltage of the second and tend to drive it as a motor in the reverse Direction.-This action retards the locomotive but is very severe on the segments of the reverse cylinder and on the motor commutators. "Plugging motors" is the application of power to the motors corresponding to the opposite direction of travel. In order to "plug" the motors it is necessary first to "buck,, them (which reverses the connections to the motors) and then to apply power. Theoretically, this should be done only in an emergency, but locomotives in mine service are frequently thus treated., Dynamic braking is retardation of the locomotive by the motors acting as generators, which take no power from the line.

Jan 1, 1942

By C. S. Allen

With increased coal production and mechanization of coal mines many transportation problems arise. The main objective is to bring the coal to the tipple or dump it as quickly as possible. Larger and faster haulage locomotives or two locomotives connected in tandem have been adopted to solve the problem. Motor equipment, gears and control systems of locomotives to be placed on these hauls are carefully studied to meet these requirements, by an analysis of motor characteristic curves with the corresponding tractive effort curves, in order to determine the correct load for maximum efficiency and to determine the approximate schedule. Control equipment is usually considered to be the controller and resistance. Sometimes accelerating curves are calculated in order to obtain smooth starting in the least possible time without slipping the wheels. Train control is very important, but why limit this to acceleration ? Deceleratien-or-retardation of the locomotive is a part of the train control and should be given as much consideration as the control for acceleration. Types or Braking Common braking systems for mine locomotives utilize screw-type mechanical brakes, dynamic braking and the air brake. For gathering and haulage locomotives, the srrew-type mechanical brake is regarded as standard. A brake of this kind cannot be applied without much effort and lapse of time. Motormen, like other people, follow the path of least resistance, and therefore have their brakes ready for instant application—which means that the shoes are virtually dragging. Even then, a motorman bucks and plugs the motors to gain time and avoid as much physical effort as possible. "Bucking of motors" consists in turning the reverse cylinder of the controller to parallel position corresponding to the opposite direction of travel while the locomotive is in motion, without applying power from an external source. When this occurs, both motors tend to act as series generators, because the armatures are revolving and there is residual magnetism in the field circuit. Owing to slight differences in motor manufacture, one of these will generate a slightly higher potential and will overcome the generated voltage of the second and tend to drive it as a motor in the reverse Direction.-This action retards the locomotive but is very severe on the segments of the reverse cylinder and on the motor commutators. "Plugging motors" is the application of power to the motors corresponding to the opposite direction of travel. In order to "plug" the motors it is necessary first to "buck,, them (which reverses the connections to the motors) and then to apply power. Theoretically, this should be done only in an emergency, but locomotives in mine service are frequently thus treated., Dynamic braking is retardation of the locomotive by the motors acting as generators, which take no power from the line.

Jan 1, 1942