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ANDERSON, ALEXANDER, Captain, Royal Garrison Artillery. ANDERSON, GEORGE K., JR., Lieutenant, 30th Co., 3d Replacement Regiment. ARMSTEAD, H. H., Captain, Quartermaster Corps, U. S. N. A. ARMSTRONG, HAROLD KERR, Student Flight Officer, U. S. Naval Reserve Flying Corps, Dunwoody Naval Training Station, Minneapolis, Minn. BAKER, HARRY A., Captain, Co. M, 53d Pioneer Infantry. BASSETT, THOMAS E., Corporal, 340th. Field Artillery, Battery A, 89th Division, A. E. F. BATES, Wm.. T., 3d Reserve Officers' Training Camp. BAUMMGARTEN, K., 1st Lieutenant, 1st Corps School A. P. 0. 703. BAYLESS, G. E. S., Supply Dept., U. S. Submarine Base, New London, Conn. BENSEL, J. A., Major, U. S. R., Board of Control, War. Construction Activities, Norfolk, Va. BILLICK, DON C., Co. C, Ordnance. Supply School. BILLINGSLEY, PAUL, 1st Lieutenant, National Army, Statistics Branch,. General Staff. BLANDY, S. H. B., B. E. F.

Jan 8, 1918

By H. M. Myers, J. E. Funk

The two-phase flow of a drilling fluid in a diamond drill bit is investigated by deriving and solving the steady-state continuity and momentum equations. Parameters in this analysis are rotational speed of the bit, local rate of addition of drill solids, non-Newtonian nature of the drilling fluid and drilling rate. The describing equations relate the flow and pressure drop to the design characteristics of the bit. The equations are integrated numerically yielding the flow-rate pressure drop characteristics for the bit, flow distribution in the bit and velocity and static pressure as a function of position in each flow channel of the bit. Results are presented for a particular bit design. This paper shows that under normal drilling conditions certain parameters may be neglected in the hydraulic calculations for a diamond drill bit. Therefore, shortened forms of the equations presented can be used to fit the proper diamond bit to the drilling fluid hydraulic system. INTRODUCTION Like the conventional bit, the diamond drill bit is a key component in the circulating system of a well. The conventional rock bit is fitted to the circulating system by selecting the flow nozzles which provide the system with the proper flow-rate pressure drop relationship. Fluid behavior across the face of the diamond bit must be better understood so the bit can also be fitted to the circulating system. Data from field-operated diamond bits are presently available and do yield some information regarding the flow-rate pressure drop relationship. However, the number of factors affecting this relationship is quite large. Drilling parameters and conditions such as bit penetration rate, bit weight and rotational speed are important. Moreover, flow channels in the bit are part of the cutting head, and introducing drilled solids into the flow channels reduces the flow area, causes additional drag and influences the pressure drop. Many tests would have to be performed if field tests were required to study all parameters which might influence fluid behavior in the bit. A logical approach to this problem is to investigate analytically the fluid behavior in the bit, including parameters that are present under normal drilling operations. This approach, which seems to offer the best method for future studies, is presented in this paper. Following is the general procedure. 1. Equations describing the most general flow channel are written and include the addition of drilling solids (which is related to penetration rate), rotation of the bit, design parameters of the bit and non-Newtonian properties of the fluid. 2. A typical bit is designed and the bit parameters are substituted into the equations. 3. Parameters are assumed (penetration rate, rotational speed of the bit, mud weight, etc.). 4. The describing equations are numerically integrated to obtain the pressure drop across the face of the bit as a function of flow rate. 5. The pressure-drop contribution of each component (friction, acceleration, etc.) is obtained from the numerical solution, and those factors which largely influence the total pressure drop are identified. Classifying the drilling fluid is important in establishing the friction factor relationship since the frictional pressure losses are the main contributors to the total pressure drop. High-density drilling fluid is commonly assumed to be a Bingham plastic. 1-3 Govier4 presented a method for selecting a friction factor in the laminar flow region from the rheology properties of the mud and a modified Reynolds number. Metzner5 presented a criterion for selecting the friction factor in the turbulent flow region. The bit design used in conjunction with the solution of the equations has divergent flow channels. Stall, which is the backflow of fluid in the channels due to eddies that occur on the outer perimeter of the flow stream, can occur in these divergent channels (or diffusers). When stall does occur the static

By John G. H. Crump, E. W. Owen, Peter P. Gregory

A noticeable decrease in activity characterized the year 1938 in the West Texas area. The total number of wells completed dropped to 2045 for 19381 as compared with 2806 completions in 1937,l a decline of 27.12 per cent. This trend contrasts the sharp increase of 76.29 per cent in completions in 1937 over the 1936 total. Of 2045 well completions during 1938,' 1788 were oil wellsll 16 were gas producers1 and 241 were listed as failures.' During 1938 the ratio of dry holes drilled to oil-producing or gas-producing wells completed was one dry hole to each 7.48 oil wells, which compares with one dry hole drilled to each 7.477 oil wells in 1937. Nearly 200 wells drilled during 1938 in West Texas were either wildcat or partly wildcat wells, an increase of about 25 over the year 1937, when approximately 175 wells drilled were so classified. The 200 exploratory wells drilled represent about 10 per cent of the total number of wells completed. West Texas produced 71,257,541 bbl.1 of oil in 1938 as compared with 75,743,000 bb1.2 of oil in 1937, a decrease of 4,485,459 bbl., or a 5.92 per cent decline. By comparison, West Texas produced 22.36 per cent more oil in 1937 than during 1936. In December 1938, the 9247 producing oil wells in West Texas1 averaged 22.08 bbl. per day per well as compared with December 1937, when 7571 wells1 produced an average of 25.18 bbl. per day per well, or a decline in daily well average of 12.32 per cent. Discoveries No outstanding new fields were discovered in West Texas during 1938, in spite of continued, active wildcatting operations. Several new oil-producing areas were, found, however, the value of which is still unknown.

Jan 1, 1939

By E. W. Owen, John G. H. Crump, Peter P. Gregory

A noticeable decrease in activity characterized the year 1938 in the West Texas area. The total number of wells completed dropped to 2045 for 19381 as compared with 2806 completions in 1937,l a decline of 27.12 per cent. This trend contrasts the sharp increase of 76.29 per cent in completions in 1937 over the 1936 total. Of 2045 well completions during 1938,' 1788 were oil wellsll 16 were gas producers1 and 241 were listed as failures.' During 1938 the ratio of dry holes drilled to oil-producing or gas-producing wells completed was one dry hole to each 7.48 oil wells, which compares with one dry hole drilled to each 7.477 oil wells in 1937. Nearly 200 wells drilled during 1938 in West Texas were either wildcat or partly wildcat wells, an increase of about 25 over the year 1937, when approximately 175 wells drilled were so classified. The 200 exploratory wells drilled represent about 10 per cent of the total number of wells completed. West Texas produced 71,257,541 bbl.1 of oil in 1938 as compared with 75,743,000 bb1.2 of oil in 1937, a decrease of 4,485,459 bbl., or a 5.92 per cent decline. By comparison, West Texas produced 22.36 per cent more oil in 1937 than during 1936. In December 1938, the 9247 producing oil wells in West Texas1 averaged 22.08 bbl. per day per well as compared with December 1937, when 7571 wells1 produced an average of 25.18 bbl. per day per well, or a decline in daily well average of 12.32 per cent. Discoveries No outstanding new fields were discovered in West Texas during 1938, in spite of continued, active wildcatting operations. Several new oil-producing areas were, found, however, the value of which is still unknown.

Jan 1, 1939

PENNSYLVANIA ANTHRACITE SECTION Executive Committee R. V. NORRIS, Chairman. CHARLES F. HUBER, Vice-Chairman, W. J. RICHARDS, Vice-Chairman, EDWIN LUDLOW, Vice-Chairman, ARTHUR H. STORRS, Vice-Chairman. CHARLES ENZIAN, Secretary-Treasurer, U. S. Bureau of Mines, Wilkes-Barre, Pa. DOUGLAS BUNTING, FRANK A. HILL, ALBERT B. JESSUP, RUFUS J. FOSTER, JOHN M. HUMPHREY, ROBERT A. QUIN. The Pennsylvania Anthracite Section held its second regular meeting at the Pottsville Club House, Pottsville, Pa., July 25,1914. The meeting was attended by more than 90 members and guests. Vice-Chairman W. J. Richards presided. Following an informal dinner, W. G. Whildin, Mining Superintendent, Lehigh Coal & Navigation Co., Lansford, Pa., presented a. paper on Steep Pitch Mining of Thick Coal Seams, illustrated by numerous drawings. This paper will be printed in the November number of the Bulletin. A lively discussion followed the reading of the paper.

Jan 10, 1914

By Malcolm C. Bridges, Ian A. Goddard

INTRODUCTION Mount Isa mine is a major producer of both copper and lead-zinc- silver ores. During the 1975/76 year, 4.3 million tonnes of copper ore were extracted at an average grade of 3.4%, and 2.4 million tonnes of lead-zinc-silver ore at average grades of 7.0%, 6.7% and 173 grams per tonnes respectively. Systematic lead-zinc mining has continued for just over 45 years and copper mining for just over 30 years. Virtually all mining is underground. The composite and somewhat schematic longitudinal section of Figure 1 shows the overall distribution of ore at the Mount Isa mine. The mine can be roughly divided into two parts about a central shaft complex: to the north there is predominantly lead-zinc ore and to the south there is copper ore. Underground operations are serviced by three shafts in the central shaft complex (P61, R62 andU62) and another to the south in the copper mining area (X41). Mining is spread over a total strike length of 2.9 km and extends to a depth of 960 m. Currently, lead-zinc ore is being extracted between 15 and 11 levels and is hauled to the central ore-hoisting shaft on 13 and 15 levels. Copper ore is being extracted between19 and 13 levels and is hauled to the central shafts on 19 level. The principal lead-zinc mining methods used now, as they have been over the last ten years, are sublevel open stoping and cut-and- fill stoping, with about equal tonnages of ore from each (Figure 2). Since about 1970, there have been several interesting developments in lead-zinc mining methods and procedures, principally: • the development of an effective method for the recovery of pillars between open stopes and • the evolution of safer and more productive cut-and- fill methods. During this period also, the focus of mining has moved 120 m deeper

Jan 1, 1977

By P. G. Coombs

The kinetics of the cyclic oxidation and reduction processes for palladium powder were examined in the temperature range 848 to 923 K. The changes in oxidation behavior due to cycling are shown to result from the decrease in surface area. Sintering is found to occur only during the reduction process. Oxidation kinetics were found to follow a cubic law with time and yield a relation for the oxidation rate constant of k x = (1.05) 106 exp P74.0). 103 / RT] The kinetics of reduction are demonstrated to be first-order with the rate constant described by the expression: kr = (7.69) 1012 exp [-(208.0) 103 / RT] . In both expressions R is in J.mo1-1.K-1.

Jan 1, 1984

By Charles Bohdanowicz

Production of crude in 1937 decreased (Tables 1 and 5) by about 2 per cent, while gas production for that period increased by about 6 per cent after a slight decrease in 1936 (0.4 per cent).' As in preceding years, drilling operations were confined principally to extensions of old oil fields in the Jaslo district (fields 1 and 7) and to some extent to such extensions in the Stanislawbw district (field 27). Deepening of old wells to horizons reached more recently by new wells gave on the average better results per well than those attained by new wells (Table 3). Reserves of proved lands have increased some by new oil fields discovered, Targowiska (field 8) and Czarna (field 19) both of which, however, have only small production. The Daszawa gas field (field 25) has not been extended, but two new gas fields were located instead— Opary, distant 25 km. (151/2 miles) northwest of Daszawa, Balicze at 12 km. (7 1/4 miles) southeast of Daszawa, and Kalusz (TESP Company exploiting potassium salts) in the vicinity of the town of that name (field 33). Drilling upon the gas area near Tarnbw and Debica, while gas-bearing sandstone had been found at depths of from 300 to 512 meters (984 to 1680 ft.) (Zdziary No. 2) and at 220 m. (722 ft.) [pressure 11 atmospheres, flow 24 cu. m. per min. (847 cu. ft.)] (Przyborbw No. l), has not as yet ascertained horizons qualified for permanent exploitation. The exploring well near Kos6w remained for the last seven months at the depth of 1420 m. because of the collapse of rotary drilling pipe. Below the 300-m. level no gas sands have been encountered at this well. Gas Reserves In connection with the broad plan of developing industrially the sector of Poland lying between the Vistula and the San Rivers, and extending farther north as far as the city of Radom, there arose the necessity of a more accurate calculation of gas reserves for the fields of

Jan 1, 1938

By C. F. Floe, N. J. Grant, A. Joukainen

A CCORDING to Guertler,¹ Smith and Hamilton were the first to study the Cu-Ti alloy system, but because of the presence of large amounts of impurities their data are inconclusive. Hensel and Larsen² in 1930 and Kroll³ in 1931 determined the high copper end of the diagram. In 1939 Laves and Wallbaum,4 in a short summary of compound determinations in titanium binary alloys, mentioned the existence of a face-centered cubic compound, Ti2Cu, with 96 atoms per unit cell, and also a compound, TiCu3, which was said to have a deformed hexagonal close-packed structure in which ordering had been observed. The existence of a compound, TiCu, was also found but the crystal structure was not determined. Craighead, Simmons, and Eastwood determined the effect of copper on the allotropic transformation in titanium. The investigation was limited to alloys containing from 0 to 5 pct Cu, using titanium of fairly high impurity content. Experimental Procedure The methods used for determination of the diagram included metallographic examination and X-ray and thermal analyses. The system was first worked out using the higher purity "Process A" sponge titanium and oxygen-free copper. The high titanium region was next determined more accurately using iodide titanium and vacuum-melted, oxygen-free copper. Preparation of Alloys: The alloys were prepared by arc melting under prepurified helium atmos- pheres in water-cooled copper crucibles. The high purity iodide titanium alloys were prepared by melting three compositions along with a specimen of unalloyed titanium in a single charge. A multiple crucible was used which consisted of a copper plate with four cup-like depressions. The unalloyed titanium was melted first and acted as a getter for any oxygen and nitrogen in the melting chamber. Since these elements harden titanium, the hardness of the standard specimen was an index of the highest probable contamination of the alloy melts. The sponge titanium alloy melts weighed from 100 to 200 g each and the iodide titanium melts from 20 to 25 g. In order to insure greater homogeneity most of the alloys were melted, turned, and remelted twice. The larger buttons made from sponge titanium were sectioned or crushed before remelting. These alloys up to 30 pct Cu were further homogenized by forging at 925°C. They forged readily up to 20 pct Cu but numerous cracks were formed in the alloys containing from 20 to 30 pct. Above 30 pct Cu the sponge titanium alloys were homogenized by heating for 200 hr at 850°C under prepurified helium. Similarly, the iodide-grade titanium alloys containing up to 3 pct Cu were homogenized at 1000°C for 24 hr and all others at 880°C for 200 hr.

Jan 1, 1953

By J. Szekely

The paper discusses simultaneous heat conduction and diffusion. These problems may arise in slag-metal kinetics and in connection with rapid heating or cooling of specimens. A mathematical formulation is given and it is shown that the transients due to either the variability of the diffusion coefficient with temperature, or thermal diffusion, are unlikely to be important for molten or solid metals and slags. The negligibility of these effects is due to the large values of the ratio thermal diffu-sivity/molecular diffusivity. In many problems relating to transport-controlled metallurgical kinetics the diffusion process is accompanied by the transfer of heat; large-scale metal refining or metal reduction may be quoted as examples. It has been shown that due to turbulent mixing in these systems the relevant transport phe- nomena are best described in terms of "surface renewal models" which require the solution of the unsteady-state diffusion equation.1-1 Problems involving simultaneous diffusion and heat conduction are also encountered in the laboratory in connection with rapid heating or cooling of specimens. The effect of variable temperature on diffusion may be twofold: a) in the absence of "cross effects" the problem, due to variable temperature, is that of diffusion with a variable diffusion coefficient; b) if the Soret Effect is nonzero, thermal diffusion must also be taken into account in the formulation. Although information is available on thermal-diffusion transients in a steady-temperature gradient,4'5 little work has been done on the effect of unsteady temperatures on either the diffusion coefficient or on thermal diffusion.' In the following a formulation is given and solutions presented with a view of examining the areas where these effects may be important.

Jan 1, 1965

By J. C. Mertz, C. H. Mathewson

Accurate knowledge of the solid solubilities of the elements that dissolve in the important base metals is needed for guidance in the preparation and heat-treatment of the alloys derived from these components. While there is still difference of opinion concerning the merits of the X-ray and microscopic methods of determining such solubilities, it is certainly true that the recent extensive use of X-ray diffraction to review the existing constitutional data has revealed many glaring inaccuracies and has brought about a much improved knowledge of phase-boundary locations, which will inevitably come to include all of the important alloy systems. When the work is done with all necessary precautions it appears that in the majority of cases where the solubilities extend beyond a few tenths of a per cent the X-ray back-reflection technique is capable of furnishing the best of information in this field. For the most prolonged and thorough studies it is obviously advisable to combine the results of several methods, in particular, to obtain an agreement between X-ray and microscopic data. This paper reviews the work previously done on the solubilities of phosphorus, arsenic, antimony and bismuth in copper and, in consideration of the uncertainties thus made apparent, presents new X-ray data obtained with the first three members of this group. Copper-phosphorus System Heyn and Baurl found that in slowly cooled alloys phosphorus forms a solid solution with copper up to about 0.175 per cent. Huntington

Jan 1, 1937

By F. B. Richards

THERE has been much interest recently in the iron ore supply of the Lake Superior district. It may be of interest to this meeting to give some thought to this situation, dealing more particularly with the present great iron ore deposits of this section. Ores from its various iron ranges are the foundation of the tremendous tonnage of steel and iron of all kinds produced in the territory along and adjacent to the Great Lakes, drawing the bulk of its supply from this northwestern ore country. For a number of years estimates have been made of the quantity of available ore. The latest study, by M. C. Lake, chief geologist of the M. A. Hanna Co., is of great interest and brings the estimated tonnage to Jan. 1. 1929.

Jan 1, 1930

Albin, B. R. Billings, Mont. '28 Amster, Nathan L 25 Broad St., New York, N. Y. '28 Bache, Jules S 42 Broadway, New York, N. Y. '28 Barrett, C. P 5338 Harper Ave., Chicago, Ill. '28 Baum, Frank G., Cons. Hydro-Electric Engr Cassel, Shasta Co., Cal. '28 Bayless, William S Russ Bldg., San Francisco, Cal. '28 Bean, Louis H., V.P., Dwight P Robinson & Co., Inc. 125 E. 46th St., New York, N. Y. '28 Beatty, Ross J 1105 First National Bank Bldg., Chicago, Ill. '28 Benwell, Oswald F 110 Gilpin St., Denver, Colo. '28 Beymer, Albert S., V P., Keystone National Bank Pittsburgh, Pa. '28 Blumenthal, Hugo 44 Pine St., New York, N. Y. '28 Boettcher, Charles Ideal Cement Bldg., Denver, Colo. '28 Bowlus, Thomas H Iola, Kans. '28 Brown, Colvin B. .. Chamber of Commerce, Washington, D. C. '28 Burt, Maxwell Struthers Southern Pines, N. C. '28

Jan 1, 1923

By J. D. Doherty

A. R. POWELL*—Mr. Doherty has outlined in a most thorough manner valid arguments for the development of an industry in this country making syn. thetic liquid fuels from coal. No thoughtful person will dispute the statement that it is essential to carry on fundamental and applied research and acquire engineering "know-how" by the erection and operation of two or three large plants in the near future. Considerations of national security indicate that this should be done, irrespective of any temporary present surplus of natural petroleum. It is the belief of the discusser and of many others interested in coal technology that our energy picture of the future will include an increasing utilization of coal directly, without any conversion to liquid fuel, as a source of energy. A fact that is sometimes forgotten is that petroleum fuels have often been cheaper than coal in many locations and over long periods of time, and this price factor has been of great importance in expanding the use of petroleum fuels at the expense of coal. Synthetic liquid fuel made from coal must necessarily be several times as costly as the coal from which it is made, and this fact alone will have a profound effect on the future energy pattern of our country. This complete change in the economic relationship between solid and liquid fuels will lead to the substitution of coal itself for many uses now supplied by the relatively cheap petroleum fuels. From the standpoint of conservation of our natural resources, the substitution of coal itself for liquid fuels wherever possible when the age of synthetic liquid fuel arrives is also logical. At least 50 pct of the energy of coal is lost during conversion to liquid fuel. Nor is this com-pensated for by a higher utilization effi-ciency of the liquid fuel in most cases. For example, modern steam-boiler plants can operate at equal fuel efficiencies with either coal or fuel oil.

Jan 1, 1950

By W. Allen

FOR the past six years gravity surveys have been used for underground prospecting in the copper mines at Bisbee, Ariz. The primary purpose of the surveys has been to reduce the diamond drilling and crosscutting necessary for exploration. Since many of the orebodies are small, and geologic control is not always apparent, any information that will direct the drilling and crosscutting is highly desirable. Because of extensive development and exploration work in the copper mines at Bisbee, it has been possible to cover more than 630,000 ft of crosscuts on 30 levels with the gravity surveys. In the process the gravity procedures have been refined to a high degree. Density Contrast: For a gravity survey to be successful, a sufficient density contrast must exist between the geologic feature sought and surrounding host rocks. Most mineralized areas will provide this contrast if fairly massive bodies are present. In the Bisbee area the entire sequence of formations, except for alluvium, appears to have specific gravities ranging from 2.65 to 2.70. These values have been determined by means of a large number of cut samples and diamond drill cores. As a further check, vertical gravity differences have been used where nonmineralized sections are known to occur.' The only known major gravity disturbances result from mineralization that has increased the density and the voids that have decreased density. The voids are caused by mining operations and by underground water movement that has developed several areas of caverns. Equipment: While not absolutely essential, a small rugged gravity meter, such as the Worden meter, is highly desirable. A tall tripod, about the height of a transit tripod, permits instrument set-ups in deep water and in locations where fallen timber and muck piles make it impossible to use a short tripod. An additional advantage of a tall tripod is that it places the meter in the center of the crosscut, reducing the error caused by the crosscut void. Size and weight are important, since the only satisfactory means of operating the meter underground is to carry it by hand. A backpack can be used in rare instances but is usually a hindrance because of the close station spacing. The operator's ability to move through tight clearances will improve survey coverage, as it is then possible to move through raises and caved areas and to pass mine cars and machinery with a minimum of trouble. Station Control: Gravity stations are normally located every 100 ft along the crosscuts, at each intersection, and in the face of all stub crosscuts. In areas of high gravity relief, or where small anomalies might be expected, stations may be located at 25 or 50-ft intervals. When possible, the stations should be offset to avoid effects of raises or other voids. The gravity stations on a level are tied to one or more base stations, which are usually located at the shaft or near the portal of an adit. The base stations may be part of a gravity control net that extends to each level in the mine as well as to the surface. Such a net extending throughout the potential area of the surveys is highly desirable, as it is then possible to compare all gravity stations on a uniform basis. The stations that are part of the base net should be carefully established by multiple readings and, if necessary, by a least squares adjustment of the loops. In some instances where levels do not have a shaft station, or where access may be blocked by caving, it may be necessary to establish secondary bases at the top and bottom of the raises that are between levels. Under fair conditions 70 to 90 gravity stations can be located and run in 6 hr by a two-man crew. The best field procedures depend on conditions. Reduction of Field Data: Most of the time required to produce a final gravity map is consumed in processing the data. Each meter reading must be corrected for a minimum of five factors that affect the gravity value in addition to the density contrast being sought. These factors are 1) instrumental drift, 2) station elevation, 3) topography, 4) latitude, and 5) regional gravity gradient. Mine openings, such as stopes and raises, will affect the value. However, it is seldom practical to make corrections for these voids. Usually a rotation is made on the field note on the station, and any

Jan 1, 1957

By B. C. Mariacher

EXTRACTION of uranium from ores is being ac-complished by processes which. for the most part, subject the entire ore to acid or carbonate leaching. Ore deposits with a U 3 O 8 content below 0.10 pct U 3 O 8 are seldom considered suitable for treatment by leaching. A preliminary concentration that would enrich the uranium content of an ore by a simple, low cost process based on physical properties of the ore might result in some low grade deposits becoming commercial ores. In addition, the process might be employed in existing operations to reduce transportation and leaching costs and to increase capacity of existing leaching plants. A study to attempt the development of a preliminary concentration process for primary uranium ores was undertaken by the Colorado School of Mines Research Foundation under sponsorship of the U.S. Atomic Energy Commission. The objective of this work was to produce concentrates containing 0.25 pct U3O8 from the low grade ores tested. Ores Tested: The main effort was devoted to the low grade primary uranium ores from northwestern Saskatchewan. Samples were obtained from the Beaverlodge operation of the Eldorado Mining & Refining Ltd. Additional primary ores, obtained from deposits in Gilpin County, Colo., contained from 0.07 to 0.10 pct U3O8. Summary of Concentration Tests: The Beaverlodge ore was tested to determine amenability of the ore to concentration by magnetic, electrostatic, gravity, and scrubbing processes. None of these produced satisfactory results. Both gravity and magnetic processes produced fairly good concentrates when closely sized fractions of the ore were treated, but on the basis of treating the total ore, recovery was poor. Preparation of sized fractions and the low capacity of equipment for suitable concentration made these methods impractical. As flotation offered the advantage of treating the total ore without intermediate sizing, the main effort was in this direction. A flotation process was developed that fulfilled the concentration objectives as set by the AEC. Pilot plant testing was used to verify results obtained from laboratory batch testing. Mineralogy: A petrographic examination of the Beaverlodge ore included a study of polished sur- faces and identification of the radioactive mineral by autoradiograph and X-ray diffraction. Approximate quantitative mineral identification was as follows: quartz, 60 pct; orthoclase feldspar, 20 pct; chlorite, 10 pct; carbonates, 5 pct; and miscellaneous minerals, 5 pct. Included in this last group were plagioclase feldspar, pyrite, mica, chalcopyrite, pyroxene, sericite, magnetite, galena, and uraninite. The most general occurrence of uraninite was in the form of crusts and thin coatings on limonite-stained grains of pyrite, quartz, and pyrite-quartz intergrowth. At least 90 pct of the uraninite was still attached to other minerals in a 100 by 200-mesh size fraction. The uraninite crusts were as small as 10 to 20 µ diam, and 5 to 10 µ thick. The Flotation Process Petrographic examinations of the Beaverlodge ore had indicated the impracticability of attempting to concentrate the uranium by floating individual grains of uraninite. Liberation of the uraninite required grinding to sizes below those suitable for flotation. However, there was preferential association of the uraninite with some minerals while others were free of uraninite attachment. The approach to the development of a flotation process was, therefore, based upon an attempt to concentrate the uraninite by floating carrier minerals. The following paragraphs discuss the various stages of the process with regard to the factors tested and the conditions under which best results were obtained. Grinding: The most effective size range for flotation was —150 mesh + 13 µ. The —13 µ material in the final concentrate had a higher U3O8 content than the total ore, but not as high as the average concentrate; however, rejection of slimes before flotation was prohibitive because of the loss in uranium carried in the —13 µ fraction. Grinding techniques which contributed to a minimum production of fines, such as stage grinding, were then employed. Quartz and Silicate Depression: These minerals represented approximately 80 pct of the ore and were free to a large degree of uraninite attachment. Significant improvement in the grade of the concentrate was obtained by depression of these minerals with hydrofluoric acid or sodium fluoride. Promoter: Selective stage flotation of uraninite carrier minerals was simplified by development of a single promoter mixture. The mixture consisted of an emulsion of a fatty acid, fuel oil, and a petroleum sulfonate and was selected after a comprehensive series of tests. It contained three parts by weight of an oleic and linoleic acid such as Emersol 300,

Jan 1, 1957

By C. A. Pierce

STUDIES of roof and floor movement are of interest to those actively engaged in mining. This is especially true in the case of an entirely new area where there is no precedent for guidance. The potash deposit being mined is in the southeast corner of New Mexico, about 20 miles east of Carlsbad. The deposit was formed by the land-locked, shallow, Permian Sea that extended through parts of western Texas, Oklahoma, Kansas, southeastern Colorado, and eastern New Mexico. The saline deposits vary from 500 to 4000 ft. thick. Neither of the companies operating in this area has delimited its com-mercial ore bodies. As all production so far has been from Federal and State leases, the Department of the Interior and the State of New Mexico have a direct interest in the development and extraction of these ore horizons. The ore is composed of an intimate mechanical mixture of sylvite (potassium chloride) and halite (sodium chloride). It lies generally in a horizontal bed varying from 5 to 12 ft. in thickness, and has a cover of approximately 1000 ft., consisting of about 500 ft. of salt beds and 500 ft, of sediments. The immediate mine roof for several feet is principally salt with some potash and clayey material. The next 500 ft. above, from the standpoint of structure, may be considered rock salt with a few inter-bedded potash members. The upper 500 ft. is of sedimentary (Permian and Triassic) origin and contains several water-bearing formations with flows varying from 10 to 500 gal. per minute. This water intensifies the interest in convergence problems as the ore and surrounding for-mations are water soluble. Potash salt (sylvite) and rock salt (halite) are plastic minerals. Plasticity is defined as a "property of solids by virtue of which they hold their shape permanently under the action of small shearing stresses but are readily deformed, worked, or molded under larger stresses."1

Jan 1, 1938

By Robert Bell

The regions to which this paper refers include the whole of the Dominion of Canada east of the 130 Rocky Mountains and north of the water-shed of the St. Lawrence. Very little exploration for economic minerals has yet been done in this immense country; but that little has already served to indicate great wealth in some localities. Having been engaged for a number of years in conducting government geological explorations in these vast territories, I shall endeavor to give the members of the Institute an idea of the prospective mineral resources of almost half a continent, as far as they are yet discernible. I purpose confining my observations to the economic minerals, and will not attempt a description of the geology of these territories; hut it will be necessary to say a few words about the rocks in order the better to explain the distribution of such minerals. The Laurentian nucleus of the continent is the principal feature of the geological map of the Dominion. It stretches from Lake Superior to Baffin's Bay, and from Great Bear Lake to the Straits of Belle Isle. Hudson's Bay itself, which is half the size of the Mediterranean Sea of the old world, lies in the center of this area. Its shores are bordered in places with newer rocks. On the west side of James' Bay (its southern prololigation) these extend inland over 200 miles, and consist of fossiliferous Devonian and Silurian strata. On the western and northwestern side of Hudson's Bay proper, altered rocks are met with, some of which resemble the gold-bearing strata of Nora Scotia, some the Huronian of Lake Huron, some the older Huronian, and others the crystalline series of the neighborhood of Sherbrooke in the Province of Quebec. Along the east coast (called the Eastmain)and among the islands lying off it, there is an interesting set of volcanic and sedimentary rocks which appears to be identical with the Animikie and the Nipigon series of Lake Superior. The area which I have characterized in a general way as Laurentian, includes tracts and belts, more or less extensive, of the Huronian series. Such areas appear to be most common, and have been best explored, in the country between the Great Lakes and Hudson's Bay.

Jan 1, 1886

By S. C. Lindsay, D. J. I. Evans

With each year that passes hydrometallurgical processes are being more widely used to recover base metals from ores and concentrates. Generally these processes involve liquid-solid separation of metal-bearing liquors from barren residues. This may be done by countercurrent decantation, thickening and filtration, or filtration or centrifuging only. Choice of method is governed by chemical and physical characteristics of the pulps; cost of labor, power, materials, and waste disposal; and even by the amount of capital available when building plans are under way. Since all these controlling factors change with time and location and with technological advances, the separation method must be re-evaluated when new plants or expansion programs are contemplated. At the Sherritt Gordon refinery in Fort Saskatchewan, Alberta, an ammonia pressure leaching process recovers nickel, copper, and cobalt from the nickel concentrate produced at Lynn Lake, Manitoba.' Pregnant liquors are separated from the barren leach residues in a thickener-filter liquid-solids circuit. The decision to use this type of separation was based on a summation of economic factors during the pilot plant and design stages. Although countercurrent decantation was not incorporated in the refinery flowsheet, the problems solved in thickening the leach residues and the experience gained in operating the commercial plant are pertinent to countercurrent decantation, particularly where it is applied to slow-settling leach slurries or slurries having a relatively high vapor pressure of a valuable, obnoxious, or poisonous vapor. The nickel sulfide concentrate is leached under pressure with air and ammonia to dissolve the nickel, copper, and cobalt and most of the sulfur, leaving a residue of iron and siliceous material to be discarded.2,3 During a subsequent distillation stage to recirculate part of the ammonia used for leaching, copper is recovered from the solution as a high-grade sulfide. After further purification to remove traces of copper and undesirable sulfur compounds, the nickel is precipitated by hydrogen under pressure at an elevated temperature. The cobalt, together with unreduced nickel, is precipitated as a mixed metal sulfide by hydrogen sulfide for eventual cobalt recovery. The end solution is then evaporated to produce a crystalline ammonium sulfate byproduct. As shown in Fig. 1, the nickel concentrate is leached in a continuous, two-stage, countercurrent circuit. In the first stage, or adjustment leach, the fresh concentrate is partially leached (60 pct) in the liquor produced in the second stage, or final leach. This system is necessary to produce a solution suitable for treatment in the copper sulfide precipitation step. Both stages are carried out at 170" to 180°F under a total pressure, with air, of 100 to 120 psig in large, horizontal, water-cooled autoclaves. The autoclaves are divided by overflow weirs into four compartments to give a certain amount of staging during leaching. In each compartment a relatively slow-speed (100 to 150 rpm) agitator maintains the nickel concentrate in suspension. Adjustment leaching, carried out in two autoclaves in parallel, is followed by a thickener-filter liquid-solids separation. Final leaching is done in two parallel trains of three autoclaves using aqueous ammonia to extract metal values from the adjustment leach residues. This too is followed by a thickener-filter circuit. Here the residues are repulped once with aqueous ammonia and twice with process water and refiltered after each repulping stage, to insure a maximum recovery of soluble metal, before being discarded to the tailings pond. The adjustment leach is followed by a seven-disk Dorr-Oliver American filter, with an agitator in the filter boot to keep the residues in suspension. Agitators have also been installed beneath the four ten-disk filters following the final leach thickener. Slurry Characteristics: A typical analysis of the Lynn Lake nickel concentrate, produced by differential flotation,' is given in Table I. The concentrate comprises 35 to 49 pct pentlandite, (Fe, Ni)S; 3 to 6 pct chalcopyrite, CuFeS,; 1 to 4 pct pyrite, FeS2; and 24 to 42 pct pyrrhotite, Fe1-2S (where x varies between 0 and 0.2) together with various siliceous minerals such as actinolite, hornblende, and talc. During leaching, the sulfur in the pentlandite, chalcopyrite, and part of the pyrrhotite is successively oxidized through the thiosulfate and poly-thionate radicals to sulfate and sulfamate." Simultaneously the nickel, copper, and cobalt are converted to soluble ammine complexes and the iron in the pentlandite, chalcopyrite, and part of the pyrrhotite is oxidized to a spongy hydrated iron oxide. The hydrated iron oxide is readily broken down by the agitation needed to sustain the concentrate in suspension and to maintain a satisfactory mass transfer of the oxygen from the air to the slurry. Table I shows typical chemical, screen, and

Jan 1, 1961

MISCELLANEOUS U.S. 0,029,598 - Extraction of vanadium values from vanadiferous iron oxide ore. A pelletized mixture of ore, limestone and bentonite is roasted in an oxidizing atmosphere to form calcium vanadates, the calcine ground, the ground material leached with aqueous ammonium carbonate or sodium carbonate to produce a vanadium-rich solution while leaving an insoluble residue of calcium carbonate and/or calcium bicarbonate along with undissolved iron values, and the solution separated and treated for its vanadium content. This patent is a reissue of U.S. Patent No. 3,853,982, dated Dec. 10, 1974. C. B. Bare and J. W. Pasquali, assigned to Bethlehem Steel Corp. Mar. 28, 1978. 7 pp. (423-68). Filed: 3-3 1-77 (783,352).

Jan 1, 1979