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By H. M. Fisher, R. E. Snow, S. C. Sun

RAPID colorimetric estimation of the amount of transparent minerals has been applied successfully to instream mill products at the cryolite flotation plant of the Pennsylvania Salt Mfg. Co., Na-trona, Pa. Apparatus used was a Bausch & Lomb monochromatic colorimeter. The optical assembly of the colorimeter, shown in Fig. 1, passes light from a 6-v, 32-cp, prefocus incandescent lamp through two condensing lenses, a heat-absorbing filter, an aperture-control diaphragm, an interference filter, and a sample container. The light finally strikes a barrier-layer photocell and the output current of the photocell is measured by a galvanometer. The engraved index line of the collective lens, located between the lamp and the galvanometer, is imaged upon the transmission scale of the colorimeter by the galvanometer mirror. Reflected light from the galvanometer mirror, the position of which is determined by the output current of the photocell, enables reading of percent transmission throughout the mineral sample being tested. The colorimeter, Fig. 1, was standardized as follows: 1—The colorimeter was' connected to a Sola constant voltage transformer, A, of 6-v output operated from a 110-v ac supply. 2—Turning the toggle switch, C, illuminated the transmission scale, B, and the control knob, D, was adjusted with a blank in the optical system until the hairline on the scale read zero. 3—A No. 430 interference filter, E, was inserted into the filter channel, F. 4—A test tube, G, containing 7 cc of distilled water was pressed into the sample space, H, and the fine adjustment knob,

Jan 1, 1955

By C. C. Popoff

The growth of the aluminum-smelting industry in the Pacific Northwest, resulting from national defense requirements, emphasized a need for local sources of alumina to supplement the alumina production shipped from the East. Because supplies of foreign bauxite were being interrupted and our domestic bauxite reserves were inadequate, it was necessary to expedite the investigation of other possible sources of alumina during World War II. A program of investigation of high-alumina clay deposits in the Northwest was carried out jointly by the Federal Geological Survey and the Federal Bureau of Mines. As part of this program the Cowlitz clay deposits near Castle Rock, Wash., were explored by the Bureau of Mines in 1942-43. These deposits are near railroads, tide-water, and Bonneville Power Administration lines and within 22 to 100 miles of 4 aluminum-reduction plants.

Jan 1, 1955

By J. D. Forrester, Erich Sarapuu, T. C. Cheasley

Electrocarbonization of coal will produce cheap fuel for electric power plants. The complete electrification of industry and domestic power consumption of gas generated as fuel in situ is a worthwhile goal for the coal industry. Production of synthetic oils, gasoline, and coal chemicals from coal gas will create additional new markets for the coal industry. RANKING high in the nation's wealth of natural resources, readily available liquid and gaseous hydrocarbons are nevertheless falling steadily in supply, while industrial requirements for hydrocarbon compounds are greater today than during the peak period of World War 11. Theoretically, several processes will increase the future supply of liquid and gaseous fuels, but the practical application of these methods depends on two controlling factors: 1—that enough carbon-bearing material be available to warrant industrial exploitation, and 2—that the given process be physically, chemically, and economically applicable. With regard to the first factor, the Western Hemisphere is fortunately well supplied, see Table I. Table I not only shows the extensive occurrence of many fuel sources, but also the wide discrepancy among quantities of fuel substances in various materials. Great quantities of potentially available fuel have not been recovered. Many coal seams and other hydrocarbon-bearing deposits are so situated that production is uneconomical by present-day mining or oilwell pumping methods, and even in many exploitable deposits where the most efficient production methods are used a large proportion of material is never recovered. Industrial production of hydrocarbon liquids and gases must be better balanced, and processes must be developed which will convert coals, oil sand, and oil shales into fabricated commodities. Among the techniques employed by scientists and engineers to secure fuel masses not economically available at present are several processes of underground gasification. These are all concerned with converting raw materials into gaseous products by thermal treatment of source substances while they are still in the ground and therefore require that a subsurface heat (fire) zone be induced to affect pervasively the buried raw fuel deposit. This is possible only where sufficiently high temperature exists and where the fuel material is satisfactorily opened or channelized. Also, to compete successfully with common mining or oilfield practices, a given method must be controllable and must not involve greater operational expense than regular methods. Preferably, it should be less expensive. Since the close of World War 11, England has continued field studies of underground gasification methods. Experiments have also been conducted in Sweden, Estonia, France, Belgium, Italy, and Germany. The U. S. Government has sponsored field studies at Gorgas, Ala., and because gasification in situ is vitally important to establishment of the nation's fuel resources, the mining engineering department at the Missouri School of Mines has been actively interested in developing the process of underground electrocarbonization. Research was conceived by Erich Sarapuu, and preliminary experiments in treating bituminous coal were begun at the school in September 1947. The work consisted of laboratory experimentation until June 1948, at which time, through a cooperative agreement with the Sinclair Coal Co. of Missouri, pilot plant studies were instituted at the Tiger mine of the Hume-Sinclair Coal Co. near Hume, Mo., to test the encouraging laboratory findings. This comprehensive field program was pursued continuously until the beginning of 1952."'" The broad term underground electrocarbonization describes a unique gasification process that releases by electricity the desired economic constituents from a raw carbon-bearing material, converting the substance of such deposits into deliverable form and greatly reducing the cost-of-mining factor. Underground electrocarbonization is not subject to the major limitations of many other underground gasifi-

Jan 1, 1955

By V. F. Parry

The major industrial uses of char or semicoke produced by low-temperature carbonization of coal4/ are as fuel for generating electric power, for blending with metallurgical coking coal as a substitute for low-volatile coals, and for gasification to produce synthesis gas for chemical purposes. This paper discusses various methods for carbonizing coal, emphasizing processing of the noncoking coals to produce char for thermal power-plant fuel and for blending with metallurgical coking coals and tar for chemical use. Low-temperature carbonization is a relatively simple processing step involving heating coal to about 900° to 1,200° F. to decompose the coal into 4 main products: (1) Semicoke or char; (2) complex, partly oxygenated liquid hydrocarbons or tar; (3) fixed gases; and (4) water. If the coal is heated rapidly under conditions that prevent decomposition of the primary products, 68 to 87 percent of the net heat in the raw coal is left in the char and 10 to 23 percent of the potential heat appears in the primary tar. Fixed gases represent approximately 7 percent of the heat and approximately 2 percent is lost or dissipated in the reactions of decomposition. Considering the different ranks of coal in the United States, the yield of tar will vary from about 9 to 44 gallons per ton, therefore it is important that each coal be carefully considered if it is to be carbonized for industrial use.

Jan 1, 1955

By R. W. Jenkins

RACKLESS equipment is being used by the T Coplay Cement Manufacturing Co. to develop a folded limestone deposit economically and safely with inexperienced men. Cost and quality of development stone results in a saving when it is used instead of outside sources of limestone. This favorable cost has been attained by following a methods development and training program described in this paper. It is expected that further savings in favor of the mined stone will be made when stopes now being prepared are in production. The present equipment will be used for loading from stope draw points. The limestone is blended with quarried cement rock to serve as raw material for Portland cement, first produced in this country by Coplay Cement Manufacturing Co. from a process developed by its president, David 0. Saylor, in 1873. For many years the cement rock alone or with a small amount of added limestone was used for raw materials. Later as specifications were raised and the richer portions of quarries were depleted, limestone from outside sources was required in increasing amounts. In 1925 D. J. Uhle, the general manager, began to prospect company property for limestone. It was known that the argillaceous cement rock was underlain by crystalline high calcium limestone. After an extensive drilling program it was found that there were large reserves of limestone underlying a recumbent syncline exposed in the south wall of the company's No. 3 cement rock quarry. A quarry was started on the vertical limb of this fold in 1927, and by 1930 the full requirement of the plant was met. After 1945 production diminished because of the occurrence of dolomite beds in the south wall. The quarry floor was then 135 ft below the cement rock level and 285 ft below surface. In 1949 Gerald I?.

Jan 1, 1955

By John L. Baum

Tests show that 50 times the earth's normal field can exist near stacked drill rods. Protection against the effect of these strong fields can be obtained by means of a removable sleeve of common iron pipe. GEOLOGISTS and engineers working around a diamond drill rig have often had the opportunity to observe the magnetic attraction of drill rods pulled out of the hole for core removal. This is sometimes an amusing pastime, but this same magnetic force inadvertently applied to a compass needle can cause it to lose its polarity entirely or to become reverse polarized when placed in the immediate vicinity of stacked rods. The first recommendation, of course, would be to avoid the rods, but in surveying an operating drillhole this is not always possible; therefore steps must be taken to neutralize the magnetic effect of the drill rods. Early in the summer of 1949 the geological department of the New Jersey Zinc Co. undertook at Franklin, N. J., an extensive program of surveying the course of deep diamond drillholes. The majority of the holes were surveyed while in the process of being drilled, and considerable difficulty was encountered. A pattern for the deviation of drillholes had been established for the district from surveys in holes from which the drill rig had been removed. It was observed that a certain operating deep drillhole being surveyed did not fit this pattern. In the resulting confusion of repeated surveys and cementing a plug for wedging operations, the lower 1800 ft of the hole was lost. Because of this incident studies were initiated to determine the source of the difficulties and to develop safeguards against recurrence of faulty surveys. Using a hand magnet easily shows that the polarity of a compass needle can be reversed if rotation of the needle is hindered through tilting of the compass or clamping of the mechanism. Such experiments are not recommended, especially in the case of the clockwork type of instruments. A magnetized hairspring in a clockwork instrument will cause the compass card to assume a periodic motion which will not stop until the clockwork runs down. Routine check readings of instrument north before and after each survey are not always sufficient to catch instances of reversed polarity of the compass needle. Although the majority of instances probably take place while the instrument is being removed from the hole with needle in fixed position, it is demonstrable that a 'luggish needle, weakened during a previous test, can be reverse polarized

Jan 1, 1955

By O. N. Carlson, H. A. Wilhelm, J. M. Dickinson

On the basis of microscopic studies, melting-point observations, and X-ray analyses, a phase diagram is proposed for the Cb-V system. A complete series of solid solutions is formed with a minimum in the solidus at 1810°C near 35 wt pct Cb. No compounds or intermediate phases were found in the system above 650°C. THERE is an ever increasing need for better structural metals and alloys for use in nuclear reactors. In addition to the normal properties of engineering structural materials, such as high temperature strength, resistance to corrosion, and fabric-abil~ty, the nuclear properties of the material must be considered. In a nuclear reactor it is important to conserve neutrons, so a material which removes these neutrons from the reaction excessively is considered to have unfavorable nuclear properties. In nuclear-reactor design the engineer must have nuclear as well as other data available on alloys in order to make a wise selection of materials. Due to the fact that many of the common structural materials have undesirable nuclear properties, it is vital that new alloys of metals having more favorable nuclear properties be investigated. Columbium and vanadium are both high melting metals, both exhibit resistance to chemical attack, and no great difficulty is encountered in fabricating them into desired shapes. With proper treatment both metals can be cold rolled extensively without failure. In addition they have desirable nuclear properties for certain types of reactors. Therefore, the alloys of columbium and vanadium should be of interest in the atomic energy program. Since an alloy-development program is enhanced by a knowledge of the phase equilibria of the components, this investigation was undertaken to establish the phase diagram for the Cb-V system. According to the Hume-Rothery rules for alloying,' the chemical similarity, crystal structure, and atomic-size factor are favorable for a complete series of solid solutions for this system. Both elements are in the same family of group V of the periodic table and thus are quite similar chemically. The crystal structures of columbium and vanadium are compatible for extensive solid solubility, since both have body-centered-cubic structures. The atomic diameters of columbium and vanadium are 2.85 and 2.62Å, respectively. This difference of slightly more than 8 pct is well within the 15 pct maximum difference allowed for extensive solid solubility. Experimental Procedures Source of Materials: Columbium powder and sheet trimmings were obtained from the Fansteel Metallurgical Corp. According to the manufacturer the metal contains less than 1 pct impurity. An analysis of the metal showed approximately 1800 ppm C in the powder while the sheet trimmings contained less than 500 ppm C. Spectrographic analysis showed minor amounts of Ca, Cr, Fe, Mn, Si, Ti, V, and Zr in both forms of the columbium. No commercial source of vanadium having the ductility and purity desired was available to the authors at the beginning of this investigation. As a result, all of the vanadium used in this study was prepared by the bomb reduction of vanadium pen-toxide with calcium employing the method reported by Long.' Yields of massive vanadium normally were about 80 pct. Chemical analysis of the vanadium prepared in this manner showed the presence of 200 to 500 ppm N and 800 to 1000 ppm C. Minor amounts of Ca, Fe, Mn, Si, Zr, Cr, and Cb were detected by spectrographic analysis. This vanadium metal was ductile and was cold rolled into 5 mil sheet. Annealing was not necessary during this rolling and the metal retained its cold-rolling characteristic after are-melting. Preparation of Alloys: The Cb-V alloys were prepared by melting pieces of vanadium sheet togethel-with columbium in the form of sheet or pellets of powder. The melting was carried out under argon in conventional arc-melting equipment employing a tungsten electrode and a water-cooled copper crucible. Each alloy was remelted three or four times, inverting the alloy after each melting in order to assure complete mixing. Alloys normally were obtained as round flat disks, weighed approximately 70 grams, and had roughly the shape of a disk 1 1/2 in. in diameter and 1/4 in. thick. Half of each alloy

Jan 1, 1955

By F. W. Wood

Ingots of zirconium and other highly reactive metals such as titanium and hafnium produced by the consumable-electrode arc-melting process require conditioning before forging or rolling. A nondestructive, arc-conditioning technique has been developed as a substitute for lathe conditioning. The technique employs en electric arc under an inert atmosphere to fuse surface metal on the sidewalls of ingots without seriously affecting the general quality of the ingots. Advantages of the practice are Increased yields, lower accumulation of inferior scrap, and actual cost savings. The work cited in this report was done on zirconium. INTRODUCTION The consumable-electrode arc-melting process as practiced at the Bureau of Mines3 4/ has proved to be a technique for producing sound, homogeneous, commercial-size ingots of zirconium and its alloys. Ingots produced by this process, however, have as a characteristic a layer of partly fused metal on the sidewall to a depth of 3/16 to 3/8 inch. To prevent extension of surface imperfections during forging, the ingots are ordinarily conditioned by removing defective material in a lathe. If the metal were less reactive or the requirements less severe, the machine chips could be recycled and remelted. However, zirconium, being so oxygen sensitive, is contaminated by merely being machined despite extensive cooling of the tool bit. Further, being contaminated with oxygen, the chip metal upon remelting is 10 to 25 Brinell-hardnese numbers harder than the ingot from which it was machined. For this reason, the ingot specifications for naval reactor use have prohibited remelting machine chips, except as required for starting pads. Again, were zirconium less costly or less in demand this ever-increasing pile could be burned to oxide and recycled through the Kroll sponge process. Economics do not favor this, as melting feed stock has an estimated increase In value over oxide of about $8.80 per pound of contained zirconium. The efforts of the Bureau to solve the problem have been directed along two lines: To reduce the contamination of the machine chips produced and to reduce the quantity of the chips produced. This paper presents the technique developed to reduce the quantity of chips produced by arc-conditioning the remelted ingot.

Jan 1, 1955

By H. H. Waller, D. L. Monroe, D. P. R. Smyth

ORIGINALLY slusher drift development was conventional, advancing the drift full 10x13-ft size at 6 ft per round. This proved dangerous and costly because the weak fractured rock of the orebody cannot be left raw for any length of time without its reacting with air to cause spalling. With this method of drift development ground conditions were worsened by many blasts and the length of time it was necessary for the drift to remain open. The problem was practically solved by driving 5x7-ft pilot drifts, then enlarging to 10x13-ft raw size by slashing with 50 to 60-ft diamond drill holes. Calculated explosive charges are used to avoid wall and roof shattering. Still more recently jacklegs and sectional steel replaced the diamond drills, resulting in the same work for less than half the cost, see Figs. 1 and 2. Crosscuts at right angles to the main haulage drifts, which cross through the center of the ore-body, divide the ore into 200-ft blocks. Eight slusher drifts on 50-ft centers draw from one block. The slusher floor is 11 ft above the rail, inclined at 1 pct grade, and the top is 18 ft below the undercut. The finished slusher drift is 102.5 or 120 ft long, depending on whether it is north or south of the crosscut. Drawpoints are staggered on each side at 17.5-ft centers. The completed drift is concreted and has an inclined ramp and a vent at the back. When crosscuts and gangways have been excavated and concreted and slusher stations installed, slusher drift development is carried out in 10 clear-cut stages: 1—Pilot drift is driven full length, 5x7 ft. 2—Station is excavated at half-way point in pilot drift. 3—Pilot drift is enlarged to slusher drift size by drilling. 4—Lower part of slash is blasted and mucked for floor concrete. 5—Floor is poured. 6—Inside upper half of slusher drift is blasted, mucked, and concreted. 7—Outside upper half of slusher drift is blasted, mucked, and concreted. 8—Drawpoints are excavated. 9—Drawpoints are concreted. 10—Concrete ramp and sheave protection recess are constructed at back end of slusher drift. A 5x7-ft pilot drift is driven full length on the center line of the proposed slusher drift using a jackleg and detachable bits on 7/8-in. hexagon steel. A 10-hp air-driven slusher and 36-in. scraper pull muck directly into cars in the main haulage crosscut, see Fig. 2. Operations are on a three-shift basis. One man mucks the previous round, drills, and blasts his round in a shift. Average advance is 5 ft per shift. When the pilot drift is completed, 7 ft of drift in the center section is enlarged to full raw slusher drift size of 10x13 ft, for the set-up to drill long slash holes. The pilot drift is slashed to full raw size from this station with jacklegs and tungsten carbide chisel bits on 7/8-in. sectional steel. Crews line the holes without trouble, working from controls established by the engineers, and the pattern of holes for the slash is arranged to give the least shattering effect

Jan 1, 1955

By A. P. Craig

"I HAVE BEEN looking forward to addressing this particular audience about Trans-Canada Pipe Lines because you men, more than most, understand the problems of the pioneer. Trans-Canada Pipe Lines qualifies as a pioneer on two counts: it is certainly pioneering a completely new industry in Canada -the long-distance transmission of natural gas -and it has at least a few of the problems that are expected and normal to any pioneering venture.This morning, I’m going to try to give you a sketchy picture of our project and a brief report on its current status and our progress to date; and then a broad picture of Canadian markets for natural gas.The physical characteristics of the Trans-Canada Pipe Lines project have been discussed so often that I am not going to take the time today to quote a lot of statistics on pipe sizes, miles, tons of steel, etc. Suffice it to say that this project, estimated to cost approximately $350,000,000, will be the longest large-diameter gas pipe-line ever built anywhere.Let me try to paint a broad picture of this great national project, using only a few basic 'round' figures.First, it involves exploration for and development of the natural gas fields in Alberta and other provinces -the cost of which will run into hundreds of millions of dollars.Then comes the gathering, stripping, and cleaning of the gas before it can be delivered to Trans-Canada. This will cost close to $100 million.Next we have the Trans-Canada Pipe Lines project for transmission of the gas to Eastern markets -estimated to cost $350,000,000 during the first three to four years and making natural gas economically available to one-half the population of Canada."

Jan 1, 1955

By Edward L. Clark, Garrett A. Muilenburg

THE first record of the discovery of iron ore in Missouri was Marquette's observation in 1673 of brown iron ore, or limonite, in the Mississippi River bluffs just north of the mouth of Apple Creek in southeastern Perry County. There is no record of any attempt to mine or smelt Missouri iron ores until 1815, when the Ashebran furnace was erected at the Stouts Creek shut-in near Arcadia, Iron County. Some limonite was treated in this furnace, the first iron furnace west of Ohio, but most of the ore was hematite from Shepherd Mountain and adjacent deposits. Eight years later, in 1823, the Perry furnace five miles north of Caledonia, Washington County, was constructed to smelt limonite ore. It

Jan 1, 1955

By Harold M. Mooney

WHEN apparent resistivity data are taken with the symmetrical Wenner 4-electrode spread, a fixed center position is used and readings are taken for values of electrode separation. Basic data consist of apparent resistivity plotted against separation of adjacent electrodes . The interpreter attempts to infer geologic structure, such as the depth to discontinuities and the nature of subsurface earth materials. An earlier paper' described methods for interpreting resistivity data. All of these involve a severe assumption, namely that the earth in the region of interest consists of horizontal layers, electrically homogeneous and isotropic. The actual earth never satisfies this assumption exactly and may deviate from it so much that none of the above methods can be applied. Attempts in three directions have been made to modify the assumption so that it approaches known geologic complexity more closely. First, curves have been calculated for dipping discontinuities. Stern' and Aldredge hose a few widely separated dip angles. Trudu confined his attention to small dips. Berel'kovskiy and Zubanov computed gradient curves for widely separated dip angles. Unz as given the most complete solution, with a brief attempt to treat the three-layer case. Second, anisotropy can be taken into account. It seems geologically probable that layered materials have different vertical and horizontal conductivities. Cagniard, Maillet, and Pirson et up methods for finding an equivalent hypothetical isotropic medium. Standard interpretation methods can be applied to this, and the actual medium can then be deduced. Belluigi'" discounts the practical importance of anisotropy; Geneslay and Rouget do not agree with his conclusion. Third, the effect of variable resistivity in a layer can be considered. Keck and Colby" examine the mathematics of an exponential increase in a surface layer. Several authors, for example, Stevenson,'' consider a continuous variation of resistivity with depth. The present paper deals with a linear variation of resistivity in a surface layer. Geologically, surface variations should be expected. Unconsolidated materials such as glacial drift show marked irregularities over short distances. The effects of weathering change with depth. The moisture content of material above the water table may vary from a dry sand to a saturated clay, and both of these will be changed by rainfall. Figs. 1 to 6 present apparent resistivity curves to show the effect of a variable surface layer. In all cases resistivity varies linearly with depth down to a depth of one unit. Material beneath this depth has constant resistivity. Electrode separation is plotted in depth units. Insets on each figure show the corresponding cross-sections, plotting true resistivity against depth. To illustrate, consider curve A of Fig. 1. The true resistivity of material at the surface of the earth is taken as 0.4 units. Resistivity increases with depth, reaching a value of 1.0 at 0.5 depth units and 1.6 at 1.0 depth units. Below a depth of 1.0, all the material has very low resistivity (zero, for purposes of calculation). Curve A in the main part of Fig. 1 shows how apparent resistivity varies for this case as the electrode spacing is increased. To illustrate further, curve E of Fig. 4 corresponds to true resistivity of 1.2 units at the surface, 1.0 at a depth of 0.5 units, 0.8 at a depth of 1.0, and 1.5 for all depths greater than 1.0. Apparent resistivity curves have been plotted logarithmically so that the shape of the curves becomes independent of the units, giving the curves wide validity. A certain drift-covered area, for example, shows a gradual decrease of resistivity from 230 ohm-meters at the surface to 150 ohm-meters at the bedrock surface, 275 ft down; bedrock resistivity is 800 ohm-meters. Curve E of Fig. 5 indicates that true resistivity decreases from 1.2 units at the surface to 0.8 at a depth of 1.0, then increases abruptly to a constant value of 4.0 units. Since resistivity and depth ratios are the same, this can be used to predict the field curve. For Fig. 5, multiply true resistivity and apparent resistivity by

Jan 1, 1955

By K. D. Jacob, W. L. Hill

Fluorine recovery in the United States has been restricted chiefly to manufacture of ordinary superphosphate and wet-process phosphoric acid. However, there is an expanding use of fluorine by industry in the form of hydrofluoric acid. Improved methods for recovering fluorine from phosphate rock would yield an adequate supply of fluosilicates and a substantial contribution to the hydrofluoric acid requirement. THE bulk of natural phosphates is comprised of calcium phosphates, which are usually apatites;' calcium aluminum phosphates such as pseudowavel-1ite; and aluminum phosphates, which occur in extensive deposits, notably on Grand Connetable Island." Nearly all deposits of current commercial value are apatite, often with admixed material of other classes. In the trade coarsely crystalline material, mainly fluorapatite of igneous origin, moves as apatite, whereas the fine-grained material of sedimentary origin is marketed as phosphate rock. The latter classification includes the phosphate shales in western North America. Although several genetic varieties of phosphate rock are recognized, most known reserves are of marine origin. Calcium phosphate is concentrated from Florida land pebble and Florida hard rock deposits and from Tennessee brown rock deposits. Phosphate Rock as a Fluorine Carrier: More than two decades ago Reynolds and Jacob surveyed the fluorine content of commercial phosphate rock from various parts of the world. Later these samples were re-analyzed by a new and more reliable method of determining flourine. The revised results with interim additions of new samples, published a few years ago in a brief summary," are shown in Figs. 1-3. Through the points representing the respective varieties of rock medial straight lines are drawn. In some instances the lines are well defined by the points; in others, where scattering renders the position of the best line uncertain, the greater weight is given points near the end of the covered range. Wherever distinct varieties are not recognized among regional deposits the custom is to identify the sample with the geographical location of the deposit. Although all rock varieties cannot be classified with certainty, the plotted results exhibit three distinct types of rock in which the F to P2O5 ratio diminishes, remains sensibly constant, and increases, respectively, as the grade of rock increases. Existence of two types of Florida land pebble, Fig. 1, is supported by additional results recently supplied by a producer." Generally speaking, phosphate rock from continental deposits carries considerably more fluorine, and apatite and insular phosphates less, in proportion to the phosphorus content than is required by the formula for fluorapatite. Exceptions to this are apatite from Canada and certain samples of Florida waste-pond phosphate. The manner in which excess fluorine is held and the nature of the carbonate in carbonate apatite are discussed in recent articles. High-grade calcium phosphates with F to P2O5 ratios less than 0.05 have been found in Curacao, Christmas Island," Mexic,13 and elsewhere. The calcium aluminum phosphates and aluminum phosphates aften carry notable amounts of fluorine. The F to P2O, ratio for one sample of pseudowavel-lite concentrate was 0.05." Aluminum phosphate from Grand Connetable Island is very low in fluorine, whereas some aluminum phosphate minerals" carry as much fluorine as highly fluorinated phosphate rock. Fluorine in Domestic Phosphate: The curves shown in the figures permit interpolation of the F to P2O5 ratio corresponding to the midpoint of selected ranges in grade of any one of the several varieties of phosphate. Multiplication of the interpolated ratio by the mid-range grade (P2O5) yields a figure for the percentage of fluorine in this grade. Then, with the use of the mid-range grade, the amount of fluorine held in phosphate reserves can be estimated, see Table I. Accordingly, on Jan. 1, 1950, the fluorine complement of known domestic reserves exceeded 420,000,000 long tons of the element. A previous figure," deduced in 1940 by a simpler procedure

Jan 1, 1955

By R. E. Collins, Paul B. Crawford

Potentiometric model studies have been made of ertically fractured reservoirs. All fractures originated at he well and extended into the reservoir for various distances. Studies were made to determine the areal sweep efficiencies for line-drive patterns. The areal sweep efficiency for the unfractured system studied war 70.6 per cent. It was found that if the vertical fracture war paralIel with the direction of flood, the areal sweep efficiency may be considerably reduced, approaching zero in some instances. If the fracture was at right angles to the direction of flood, the areal sweep efficiency may be greater than that for the unfractured system. The sweep efficiency depends on the length and orientation of the fracture and direction of the flood. It is concluded that considerable effort should be made to determine the nature and direction of fractures before initiating a flooding program. Such information is required if the maximum sweep efficiency is to be obtained and channeling is to be reduced to a minimum. INTRODUCTION The recent development and application of fracturing techniques to petroleum reservoirs has served to focus considerable attention on the effect such fractures may have on secondary oil recoveries. The opinion has been expressed that fracturing may not alter appreciably the water flooding possibilities. Others have thought that channeling may occur and that the creation of fractures may serve to substantially reduce the possibilities of conducting a successful water flood in the reservoir The purpose of this study was to obtain quantitative estimates of the effect of vertical fractures on the areal sweep efficiencies of line-drive flooding patterns. In presenting this work on vertical fractures, it is not to be inferred that the present fracturing techniques result in producing a vertical fracture. The types of fracture which are created are not known with certainty in every case. Some are believed to be horizontal, others inclined along bedding planes, and some data indicate vertical fracturing. Each fracture will probably exert a different influence on the sweep efficiency of the secondary recovery program; consequently each fractured well should perhaps be considered as a separate problem, requiring a careful analysis. Preliminary work on the effect of fractures on producing capacity of a well indicated that for practical purposes many fractures may be treated as having a nearly infinite permeability This assumption appears to be valid in some instances where the permeability of the fracture is only 10 or 100 times greater than the permeability of the matrix. It is known to depend on the type, orientation and length of the fracture. In realization that fractures of only 0.01 in. width have calculated permeabilities near 5,000 darcys, it is believed that many fractures may be considered as of near infinite permeability without incurring serious error. Primarily for this reason this study was made with fracture analogs of very great permeability compared to the matrix. FRACTURED SYSTEM ANALOG The pattern sweep efficiencies reported here were obtained with the aid of a potentiometric model. The application of this type model for solving reservoir problems has been described in the literature.1,2,3 The electrolyte reservoir was 20 in. wide by 30 in. long by approximately % in. deep. The liquid depth was constant, which corresponded to a constant permeability-feet product from the injection to the production well. The electrolyte reservoir was built to correspond to a unit in the line-drive pattern shown in Fig. 1. This figure shows part of an infinite array of wells having alternate rows of fractured and unfractured wells. The area within the dashed rectangle shows the area generally studied in the model. In using the model to represent a fractured system, a thin strip of copper was soldered to the well or wells and the length and orientation of the strip was made to conform to the vertical fracture under consideration.

Jan 1, 1955

By J. H. Rushton, L. H. Mahony

Principles of fluid motion and turbulence which have been found to be of use in mixing and agitation problems are discussed, as well as suggested applications in extractive-metallurgy processes. Various types of impellers are described, together with other conditions that affect flow pattern and turbulence. The choice of equipment for particular requirements is considered, and equations for power input are given. Modern heavy-duty mixing and agitation equipment can take an increasingly important part in such extractive-metallurgy processes as solids handling, crystallizing and leaching, chemical operations, and flotation. Application of mixing and fluid-mechanics principles to extraction methods can lead to greater process rates and a resultant saving in time and money. MIXERS are being applied with increasing frequency to problems in the metallurgical industries. The increase represents in part the modernization of mechanical equipment used through the application of unitized drives, modern electric motors and speed reducers, and recently developed materials of construction. Some applications have resulted from process changes and the use of techniques for operations similar to those that have been demonstrated and proven in the chemical industry. Still more applications are being developed through the use of fluid-dynamics principles not previously recognized as applicable to these operations. It is the purpose of this paper to describe those principles of fluid motion and turbulence which have been found to be of use in mixing and agitation problems, and to suggest applications for them in the extractive-metallurgy processes. Modern mixing equipment has been designed to provide the fluid motion consistent with that best suited to different operations.' When properly applied, these more efficient fluid-handling mixers result in lower process-production costs. Metallurgical operations usually involve large amounts of solids. The characteristics of the ore control the process and only enough liquid is added to provide sufficient fluidity for handling and processing. The reactive part of the ore, the mineral value to be extracted, is likely to be low in percentage, and the quantity of reagent required is similarly low, but excessive dilution makes additional reagent necessary and therefore is undesirable. Thus, the primary problems in minerals processing are to handle as high a concentration of solids as possible and to distribute reagents uniformly. These pulps, or slurries, are in a range of solids concentration where hindered settling rates apply, and care must be taken to agitate and mix sufficiently to maintain suspension and flow throughout the equipment. Such slurries have high densities which are the weighted average of the components and viscosities that are spoken of as "effective viscosities" but are difficult to evaluate. It is convenient to separate metallurgical operations involving mixers and agitators into three categories: solids handling, mechanical or physical operations, and chemical operations. In each of these categories fluid motion is required to maintain suspension, to distribute surface-active agents, to blend components and reactants, and to induce high rates of chemical reaction. After a discussion of some basic principles of mixing and agitation, examples of application to the three categories will follow. Fluid Motion The fundamental problems of mixing and agitation of liquids have to do with the mechanics of fluid streams and the means by which they are

Jan 1, 1955

By E. N. Gilje

"IntroductionDURING the past five years, over two miles of shaft sinking have been completed at Falconbridge. As sinking has continued, we have improved our equipment and technique with a resulting increase in efficiency. Throughout this period, all sinking has been done by contracting firms, under our direction and to our equipment specifications, with highly satisfactory results. Since 1949, Riddell shaft mucking machines have been successfully used in conjunction with the •benching method of sinking. Our most recent practice can best be illustrated by describing the method used in sinking No. 1 shaft at the Fecunis Lake mine.No. 1 Shaft, Fecunis Lake MineNo. 1 shaft, consisting of six compartments and measuring 13 ft. 11 l/2 in. by 19 ft. 9 3/8 in., outside timber (Figure 1), will be sunk to a depth of 4,170 feet. It was designed to accommodate two 10-ft.diam. Koepe hoists."

Jan 1, 1955

By O. Cuscoleca

The Austrian plants at Donawitz and Linz were the first to blow steel with high purity oxygen. The paper shows how the process was developed and gives a survey of the results achieved to date. Compared with open hearth steel of equal quality, the obvious advantages of oxygen steel are the low investment and conversion costs. The nitrogen content of oxygen steel is not higher than that of good open hearth steel. Rimmed and killed steel products made by the oxygen process are superior to rimmed and killed steel products made by the best open hearth process. TWO new steel plants using the principle of making steel with high purity oxygen went into operation in Austria, one at Linz (Vereinigte oster-reichische Stahlwerke) in November 1952 and the other at Donawitz (dsterreichisch-Alpine Montan-gesellschaft) in May 1953. An estimated 330,000 metric tons were produced by the two plants in 1953, the first year of operation. In 1954 production by Austrian plants- using the high purity oxygen approach will be raised to 450,000 metric tons. This will be in addition to facilities in Huckingen (Man-nesmann A. G., Düsseldorf, Germany), those planned and under construction at Hamilton, Ontario, Canada (Dominion Foundries and Steel, Ltd.), and elsewhere. It is expected that by 1955 a total capacity of one million tons of high purity oxygen steel will be available. At the close of the last century, because of increasing availability of scrap, open hearth steel production forged ahead at an ever increasing pace. This caused world steel production to rise faster than pig iron production and finally to overtake it. In 1911 65 million tons of steel were produced against 64 million tons of pig iron. World steel production was 209 million tons in 1951, while pig iron output was only 147 million tons. The 1951 breakdown for steel shows 118 million tons made from pig iron, while the balance, 91 million tons or 44 pct of the total tonnage of steel produced, was made from scrap. This increase in crude steel production on a basis of high scrap consumption led to the world market scrap shortage and soaring scrap prices. In the United States and other nations, production of converter steel has fallen far behind that of open hearth steel. The surplus of processing scrap, un-absorbable by converting mills and thus a welcome source for open hearth and electric melting shops in the past, together with regular scrap purchases, no longer can meet requirements of both open hearth and electric steel production in Europe. In addition, recent modification of the basic bessemer process, blowing with oxygenated blast, substantially raised the scrap capacity of these plants. Any further expansion of world steel production will necessitate utilization of self-containing processes in regard to scrap because of the shortage.

Jan 1, 1955

[Andrew Allen Glenwood Springs, Colorado C. S. Allen 211 Utah Oil Building, 10 West Broadway, Salt Lake City 1, Utah Earl W. Allen 257 Rio Grande Street, Salt Lake City, Utah Louis B. Allen 411 Walker Bank Building, Salt Lake City, Utah Thomas Allen State Coal Mine Inspector of Colorado, Argonaut Hotel, Denver, Colorado J. B. Ambler P. O. Box 591, Denver, Colorado Nephi W. Anderson Dragerton, Utah F. F. Argust 270 High Street, Pueblo, Colorado E. K. Arnston : 4601 West 26th Avenue, Denver 12, Colorado Abel Audin Somerset, Colorado Clement Audin, Jr Paonia, Colorado Ralph Andin P. 0. Box 285, Somerset, Colorado ]

Jan 1, 1955

By J. Offeringa, C. van der Poel

The report describes scaled model experiments on the recovery of oil from porous sands by the injection of miscible liquids [solvents] and the subsequent recovery of the solvents by water flooding. Special attention has been paid to the scaling problem. Scaling rules are deduced and their correctness is checked by carrying out experiments in tubes of various sizes. Data are submitted on the effect of the viscosity ratio oil-solvent on the efficiency of the displacement process, the viscosity ratio varying from 1.5 to 5,000. Breakthrough recoveries obtained when flooding with kerosene as a solvent, are equal to those obtained in water flooding at the same viscosity ratio, indicating that he phenomenon of viscous fingering is decisive. After breakthrough, however, flooding with a miscible liquid becomes much more efficient than a water flood and this is more pronounced as the viscosity ratio oil-solvent is higher. Although high oil recoveries are obtained and most of the solvent in the sand can be recovered by a subsequent waterdrive, a straight flood with miscible liquids does not appear economically attractive because of the large quantities of solvent required. Therefore, a circulation process is introduced, the features of which are described in this paper. Though by means of this circulation process the amount of solvent required can be drastically reduced, the process is not yet economic. SCOPE OF THE INVESTIGATION The recovery efficiency by water flood of oils of viscosities exceeding 50 cp is low. As interest in production of these highly viscous oils is growing at the moment, it was considered worthwhile to gain basic information on the feasibility of first flooding such oils with a low viscous miscible liquid [solvent] and subsequently producing the latter in a conventional way, e.g. by water flooding. The method has the obvious drawbacks that return on the investment is delayed until water flooding has begun and that it requires the investment of a more valuable product to produce a less valuable one. On the other hand, it may be expected that the displacement of one liquid by another is more efficient when the two liquids are miscible than when they are immiscible. As laboratory data on this point are scanty, the present investigation seemed justified. Another argument for carrying out these experiments is that they present a valuable contribution to our concepts of the scaling problem in reservoir engineering studies. Some major difficulties which are encountered in the scaling of the waterflooding process such as the scaling of the capillary forces and the difficulty of obtaining the same permeability characteristics in the model do not present themselves in this case. As the fluids are miscible capillary forces do not play a role and as there is only one fluid phase, Darcy's law is valid in its original form. Consequently it is felt that the scaling rules can be deduced with a greater amount of certainty. An essential feature of the experiments is that they are scaled model ones, which are, however, somewhat

Jan 1, 1955

By F. D. Rosi

The plastic properties of extended silver and copper crystals of varying purity were studied as a function of crystal orientation in the early stages of flow. Variations in the gross shape of the shear stress-shear diagram and in the properties of critical Variationsshear stress and shear-hardening coefficient were correlated with changes in slip-band developments. The phenomenon of work hardening is discussed in terms of the existing dislocation theory. IT appears certain from the early studies on the deformation characteristics in metal crystals1-3 that plastic flow takes place in a preferred crystallo-graphic slip system which is determined by the geometry of the crystal (law of maximum shear stress), and that the value of shear stress for this system is independent of crystal orientation (law of critical shear stress). Experimentally, it has been demonstrated further that the law of critical shear stress can be extended to include extensive plastic flow.&apos; Thus, the shear-hardening coefficient, which is defined as the slope of the shear stress-shear curve, also is considered to be independent of orientation. It is noteworthy that deviations from this empirical shear-hardening law have been reported in cubic crystals whose initial orientation favors slip on more than two systems.&apos; Moreover, this law appears to have been derived from stress-strain data relating to relatively high values of shear strain (0.5 to 4), where widespread slip and complex distortions can be expected, regardless of crystal orientation. Since existing data indicate that the strain inhomogeneity in a crystal is manifested particularly in the early stages of flow, it would appear that a more exacting test as to the fundamental nature of the shear-hardening law would be to investigate systematically the shape of the stress-strain curve as a function of crystal orientation in the earlier stages of deformation. With regard to the general form of the shear stress-shear curve for cubic crystals, early studies show a parabolic hardening law, where the shear 1 stress is proportional to the square root of the shear-strain. Since this law was predicted theoretically by Taylor" in his original dislocation model for hard- ening, it has been accepted widely as a fundamental flow characteristic of cubic metals. However, it was recently pointed out by Masing6 hat the parabolic law is not the elementary form of hardening in cubic crystals, but instead is a consequence of complexities in the flow process (e. g., deformation bands and unpredicted secondary slip). Thus, in the very early stages of deformation (< 2 pct extension) where the occurrence of such complexities is unlikely, a different strain-hardening behavior might be anticipated. This view is supported by the recent work of Rosi and Mathewson7 on high purity aluminum where a linear law was obtained for extensions up to 2 pct. A linear hardening over a wider range of deformation also was reported by Rohm and Kochendorfer8 ho deformed aluminum crystals under conditions approximating pure shear. Even more pertinent is the recent evidence of Masing and Raffelsieper9 on high purity aluminum crystals. It was found that for crystals having initial orientations near a <l00> or <1ll> axis, a high hardening was obtained, whereas crystals with a <110> orientation exhibited a low and linear hardening curve followed by a region of more rapid hardening. Since much still remains obscure concerning the details of strain hardening as well as slip-band formation in face-centered-cubic crystals in the early stages of deformation, the present study was undertaken to evaluate the effect of crystal orientation on these two important manifestations of glide. It is important to note here that at the time of this study, Lucke and Lange11 presented information on the orientation-dependence of the shape of the strain-hardening curve in aluminum of various purities. Their work, which essentially represents an extension of that of Masing and Raffelsieper, in many respects is parallel to that of the present study. Experimental Procedure Production of Single Crystals: Single crystals of silver and copper of varying purity were used in

Jan 1, 1955