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By Henry J. Welge

Laboortory apparatus has been devised which permits study of the displacement of oil from cores by water and by gas. The cores used contained interstitial brine as well as oil. Experiments were run to determine the comparative effect of varying the properties of the fluids used. No great effect was noted on the maximum displacement achieved. This observation made it unnecessary in initial work to use fluids in their exact reservoir conditions. Consequently, the displacements were run at near-atmospheric pressure in Pyrex glass equipment, using stripped crude oils. Introduction and Theory The chief object of this work has been to determine the efficiency of gas and water as primary agents for displacing oil from reservoir rock under laboratory conditions in which capillary phenomena were predominant. To this end the maximum displacement of oil from cores has been ascertained. This maximum displacement may not be equal to the maximum displacement from a reservoir; but it will be a close approxirnation to it sometimes, and other times the laboratory information will be useful in reservoir engineering predictions. It is] believed that the laboratory experimental maximum represents the upper limit for the reservoir recovery. The experiments were carried out by obtaining cores of interest from the reservoir, and filling the pores with interstitial brine and oil with the restored state tech- nique.1 Then the oil was displaced from the core as described later, either by brine from below, or by gas from above. The former type of displacement suggests analogy to production by water drive, but not to water flooding, for reasons discussed below. The latter type of displacement is believed to simulate production by gas cap displacement. The displacements were performed by what may be termed the capillary-pressure method. The cores are placed in capillary contact with an oil-wetted membrane which has very small pores (about I micron in diameter). Pores of this size will transmit oil but Prevent the Passage of gas or water, unless the Pressures used are higher than the capillary Pressures employed in this work. Accordingly, use of the membrane makes it possible to apply a capillary pressure differential between the displacing phase and the oil in the core. AS in the displacement of brine toward that condition of oil and water distribution existing originally in the reservoir,' it may not always be desirable to reduce the oil saturation in this experiment to an absolute minimum. With reservoirs having small productive closure, it is preferable to evaluate and use a displacing pressure no higher than the capillary pressure, Ap, which will ultimately become available for displacement in the reservoir. This pressure may be calculated from a knowledge of the density of the oil-in-place, do, the density of the displacing phase, dl, and Ah, the vertical distance between the section of the

Jan 1, 1949

By G. M. Mason

History Due to its strong affinity for oxygen, aluminium as an element was isolated only after many other rarer metals had been prepared and studied. It was just over one hundred years ago that two scientists, Hans Christian Oersted and Friedrich Wohler, produced the first few pellets of aluminium and determined its physical properties. These early investigators used metallic potassium as a reducing agent. Later, St. Claire Deville, by using sodium instead of potassium, improved the yield in the reduction and produced some pounds of aluminium. From a figure of about $110 per pound in 1855, the price had fallen to about $4 per pound by 1886. In that year, a very important discovery was made independently by Charles Martin Hall in Oberlin , Ohio, and Paul . Heroult in France. Stated in its simplest terms, the discovery of Hall and Heroult was that aluminium could be produced by electrolyzing a solution of the oxide in molten cryolite. The application of the Hall-Heroult process made the metal immediately available in commercial quantities and, by 1900, the price had fallen to 23 cents per pound. The growth of the industry in the past fifty years is well known, and is outside the scope of this paper. It is, perhaps, not generally known that aluminium was first produced in Canada at Shawinigan Falls in 1901. Since that time, production has been continuous at one or more smelters in the Province of Quebec. The largest of these smelters, which is also the largest aluminium smelter in the world, is at Arvida, Quebec, on the Saguenay river. A large power potential and access to ocean shipping were the two principal reasons for the establishment of aluminium production facilities in the Saguenay district. It is interesting to note that the aluminium industry, as a large user of power, took an active part in the early industrial development at Niagara Falls and Shawinigan Falls, and is now repeating the process in the Saguenay valley.

Jan 1, 1949

By R. V. Higgins

The, large volumes of low A. P. I. gravity oil that will remain in many shallow fields of California when such oils no longer can be recovered profitably by pumping are important potential sources of oil when ways and means are found to remove them from the reservoir sands. Estimated ultimate oil recoveries in many of the "low-gravity fields" of the San Joaquin Valley in California indicate that less than 25 percent (on the average) of the original oil will be recovered. These oils have viscosities of 1,000 or more centipoises at reservoir temperatures and, because of their high viscosity and low intensity of the reservoir pressure, the quantities that will be retained in the sands ultimately are so large and their recovery is such a challenge to petroleum engineers that the Bureau of Mines has made the study of their possible recovery one of its major projects. The fact that a slight increase of temperature materially reduces the viscosity of low A. P. I. gravity oils has 'prompted the use of various means for increasing the temperature of the oil in the reservoir sands and hence the rate of flow and Completeness of recovery of the oil. Hot oil, hot water, heated air, and natural gas, steam, super-heated steam, and the products of combustion have been circulated in wells to raise the temperature of the oil it the sands; electric heaters have been installed in wells, and butane and air have been intermittently exploded and the flames propagated through the tubing to the "pay" zones. None of the methods, however, have been entirely satisfactory for the recovery of large quantities of oil.

Jan 1, 1949

By Walter Rose

In a recent technical note, Owen Thornton&apos; suggests that wetting liquid relative permeability may be derived from the relationship: where Pd/Pc is the ratio of displacement pressure to capillary pressure at the wetting liquid satdration Sw, and I is the resistivity index at this saturation. Thornton shows that this expression gives values for relative permeability in good agreement with those experimentally determined by Leverett. We have recently developed another expression for krw which seems preferable to Equation 1 since it requires fewer experimental data for its verification. This equation may be derived in the following manner. Making use of the analogy between mass transfer of fluid in a porous medium and electrical conductivity through the fluid in the same medium, we can postulate a hydraulic formation resistivity factor analogous to an electrical formation resistivity factor. From Poiseuille&apos;s Law the resistivity to flow in a tube of radius R is of the form 8u/R², where is the viscosity of the fluid. Similarly7 the resistivity to flow in a porous medium of the same dimensions as the tube is given by D&apos;Arcy&apos;s Law as where k is the permeability of the medium. The hydraulic formation resistivity factor, Ff, is thus the quotient of these two resistivities, or: But permeability is defined by the KO-zeny equation as: where: r is the average pore radius of the porous medium, L. is the average tortuous length of the average pore, + is the porosity, and L is the bed length. This gives Fr as a function of the (La/L) .-.-.tortuosity ratio and porosity, or in explicit form: where: T = (La/L)&apos;, and N is the number of pores of radius, r, which will be found in any cross sectional area, x R&apos;. That is, N = R²/r². It will be clear from the foregoing considerations that the hydraulic formation fat. tor should be dependent on the value of R which in every case must be arbitrarily selected. By analogy, the hydrauiic formation factor, Fcf, characterizing the porous medium at Sw<1,will be given by: R² NT. where ke is the effective wetting liquid permeability obtaining at Sw<l, and Te is the square of the (Las/L) ratio defined by Thornton. Therefore, the hydraulic resistivity index, If, is: However, Thrnton gives I as: and it follows that: Equation 2, when checked against the data of wyckoff and Botset² and some of the Morse et al data³, gives comparisons as shown at the bottom of this page. Computations based on Leverett&apos;s data, quoted by Thorntaon, gives computed krw values somewhat lower than the experimental values. Moreover, other instances can be cited where we find that neither Equation 1 nor Equa tion 2 is checked by experimental data. The work of Botset&apos; on the Nichols Buff sandstone and Morse et a1 on oil wetted Bradford sandstone are such instances. A reason for these discrepancies may lie in the fact that if both Equations 1 and 2 are presumed always to give accurate values for krw, it follows that PC = Pd/Sw". Such an expression, it is well known, is too simple to describe accurately the capillary pressure behavior of all porous media. However, it is believed that a particular advantage resides on the use of Equation 2, since it is not dependent for its utility on a knowledge of the capillary pressures obtaining in dynamic flow svsterns. No practical technique for measuring the capillary pressures characterizing the fluid distributions in dynamic flow systems has yet been proposed. Acknowledgment is given to Dr. Paul D. Foote, executive vice-president of the Gulf Research and Development Company, for permission to publish this note. 1. Owen F. Thornton, "A Note on the Valuation of Relative Permeability," J. Pet. Tech., 1:7, Section 1, July, 1949. 2. R. D. Wyckoff arid H. G. Botset, "The Flow of Gas-Liquid Mixtures through Porous Media," Physics, 7:325-345, 1936. 3. R. A. Morse, P: L. Terwilliger, .and S. T. Yuster, "Relative Permeability Measurements of Small Core Samples," Oil and Gas J., 46:109-125, Aug. 23, 1947. 4. H. G. Botset, "Flow of Gas-Liquid Mixtures Through Consolidated Sand," Trans. AIME, 136:91, 1940.

Jan 1, 1949

From 1886, the first year shipments were recorded, until 1939, the Leadville district produced over 3,500,000 long tons of manganiferous ore. Of this total, metallurgical-type ores comprised 27 percent and manganiferous iron ores suitable for fluxing in lead smelters made up 73 percent. In 1939, the Bureau of Mines, as part of the strategic minerals investigative program being conducted at the time, requested the Mineral Industries Survey Section of the Mining Division to make a comprehensive study of the manganese resources of the Leadville area. This investigation, owing to the caved and flooded condition of the underground mine workings, was necessarily confined to a study of old maps, production records, and geological reports, and the results were published as an Information Circular in July 1940.

Jan 1, 1949

By R. W. Geehan

The Yellow Pine zinc-lead mine, situated about 3-1/2 miles northwest of Goodsprings, Clark County, Nev., was explored by core drilling during the fall of 1942 and the summer and fall of 1944 by the Bureau of Mines in cooperation with the Federal Geological Survey. The Yellow Pine district was located in 1855. The original claims covering the Yellow Pine mine were located in the following years: Rover, 1892; Hilo, 1897; Bybee, 1900; Radio and Como, 1906. The mine was operated by the Yellow Pine Mining Co. from 1906 to 1929. This company was dissolved in 1935. At a later date, S. E. Yount and the Security First National Bank of Los Angeles, Calif., acquired ownership of the mine. Since 1939, the mine has been operated in a small way by leasers. Production has been at least 175,000 tons averaging about 10 percent lead and 25 percent zinc, with a gross yield of $5,785,174. Virtually all ore produced by this mine consisted chiefly of oxidized zinc and lead minerals and came from long, pipelike shoots in the 100-foot thick Yellow Pine limestone. Ore shoots are sharply defined; elsewhere, the limestone is nearly barren. Numerous faults cut the mine into blocks. Ore is now mined in an open stope, but in the past, large tonnages wore mined in stopes by square-set methods. Ore produced recently was shipped direct to smelters unprocessed; in the past, it was milled by gravity and produced separate lead and zinc concentrates.

Jan 1, 1949

By R. A. Stauffer

The chief ore of lithium is spodumene, a lithium-aluminum silicate containing up to 3 pct lithium. The preparation of lithium salts from spodumene is costly because of the low concentration of the metal and the difficulty of separating it from the other alkali metals present. Dr. Kroll&apos;s and Dr. Schlecten&apos;s recent paper1 has reviewed various methods proposed for preparation of lithium. It will be seen from this paper that the methods proposed to date have required one or another of the lithium salts as a raw material. The high cost of these salts has been the chief item of expense in the lithium metal produced. This situation has made it interesting to investigate the possibility of producing the metal directly from spodumene. When spodumene is calcined at 1000ºC in the presence of lime, lithium oxide is replaced in the ore as is evidenced by the fact that it may be leached from the reaction mixture with water.2 Free energy calculations suggested that lithium oxide might be reduced with aluminum or silicon and the lithium produced distilled from the reaction mixture by heating-to about 1000ºC in a vacuum. These facts taken together suggested that it might be possible to prepare lithium metal directly from spodumene by heating a mixture of the ore, lime, and a suitable reducing agent in vacuum and distilling off the lithium metal as formed. The Replacement of Li2O in Spodumene A series of experiments were undertaken to determine the ratio of lime to spodumene necessary to free substantially all the lithium oxide Present in the Ore. The spodumene used in these experiments was supplied by Foote Mineral Co. in the form of powder of about 200 mesh&apos; Analysis of the spodumene was as follows: PCT SiO2......................... 62.7 Al2O3........................ 29.9 Li2o......................... 4.9 Na2O........................ 0.89 CaO......................... 0.75 MgO........................ 0.49 K2O......................... 0.44 Lime (cp —200 mesh) was added in various ratios to spodumene and thoroughly mixed in a ball mill. Two hundred gram samples of spodumene-lime mixtures were charged loosely in fireclay crucibles and held at 1000ºC in a gas-fired furnace for 5 hrs. The product in each case was ground to pass 140 mesh and then placed in a thimble in a soxhlet extractor. Extraction with water was carried on for 6 hrs. After this period the extract was reduced in volume by evaporation in order to precipitate dissolved lime and was then filtered and titrated with 1/4 normal hydrochloric acid. After correcting for residual calcium oxide in solution, the alkali present was calculated as lithium oxide. The results of these experiments are shown in Table I and Fig I. At lime-spodumene ratios less than 1.5, the products were invariably fused. At the ratio of 1.5 and above, the product was loosely sintered with little evidence of fusion. When the ratio of lime to spodumene was three, replacement of alkali oxides in the

Jan 1, 1949

By Wm. D. Owsley, R. E. Moeller

In the petroleum industry, the process known as oil well cementing is the equivalent of pressure grouting in the mining and construction industries. The science of oil well cementing has been known and practiced for many years. At the present time it has been brought to a high degree of perfection in the handling of many different types of jobs, all of which are known as oil well cementing. In order to understand more easily these processes, we will discuss briefly several different types of oil well cementing jobs, the purposes for which they are performed, and the methods of their accomplishment. It is believed that some of .these procedures, in a modified form, can be applied to pressure grouting projects in general. Originally, oil well cementing consisted of placing a neat cement slurry about a steel casing set in a bore hole. Cement is used about these casings to protect them from the corrosive effects of ground waters, to protect the oil and gas zones from encroachment of undesired water from other levels, to prevent the blowing out of high pressure oil and gas zones from shallow levels, and to strengthen generally and support the casing being placed in the bore hole. After the bore hole is completed to the desired depth, casing is run into the well, pumps are attached to this casing and circulation of the mud fluid in the hole is established; this circulation being from the inside of the casing into the annulus about the casing, and thence back to the surface. Next, a cement slurry is mixed and pumped into the inside of the casing string, A plug, which acts as a piston within the casing, is placed above the cement slurry. This is to separate the top of the cement slurry from mud or water which is used to shove the cement down through the casing and out into the annulus. This plug further serves the purpose of shutting off against a baffle in the bottom end of the casing string, thus stopping the flow of cement out of the casing and leaving the shoe, or lower end of the casing, surrounded by cement and also filling the annulus there above. Casing strings cemented in the above described manner may vary in length from a few feet of large diameter pipe to very long lengths of smaller pipe. As of the present date the longest string of casing which has yet been cemented is some 16,600 ft deep. As a general rule, the deeper the well and the more cement slurry used, the higher the pressures which are required to place the slurry. In parts of West Texas, it is not uncommon to use 4000 sacks to cement a single string of casing in place. Machinery used to mix the cement and place it into the casing is mounted on heavy duty, powerfully equipped trucks which are capable of negotiating any type of highway or country roads. These trucks

Jan 1, 1949

By J. Tuzo Wilson

Abstract A division of part of the Canadian Shield into provinces believed to represent former mountain belts is attempted, special emphasis- being laid upon radioactive age determinations, the relation of Archaean to Proterozoic rocks, and the structure of the Grenville province. Original data on the distribution of radioactivity within several batholiths and their connection with ore deposits are&apos; presented. The possible growth of continents and a theory which explains the shape and distribution of the world&apos;s younger mountain and island areas are discussed and applied to the Canadian Shield in an attempt to explain how the growth and present structure came about. The relation of the major mining camps of northeastern Ontario to fault zones is discussed in the light of this theory. Introduction The valuable symposium issued by this Institute to celebrate its jubilees emphasizes the important role of major faults in determining the location of many ore deposits in the Canadian Shield. Many of these faults are known to extend for long distances. It is therefore widely appreciated that it would be of great value if the pattern, ages, and causes of these faults were better understood. This would be particularly useful if the relations of the major fault systems to petrological and metallogenic provinces and to former mountain belts could be established. Each year the volume of available data, and hence the chance of reaching a fuller solution of this problem, increases. This article is a review of the present situation. It is accordingly based almost wholly upon the work and ideas of others, which are gratefully acknowledged. Naturally, responsibility for the selection and marshalling of the data is the writer&apos;s.

Jan 1, 1949

By Benjamin N. Webber

The peninsula of Nicoya contains virtually all of the known manganese deposits of Costa Rica. These are south and west of the Tempisque River, which flows across the peninsula near its juncture with the mainland. This peninsula forms the westernmost part of the province of Guana-caste, which is the northern. of the two Pacific Coast provinces of Costa Rica. The larger villages are connected by ox-cart roads and trails. The area may be covered on horseback without difficulty except during the rainy season, when rivers are in flood and difficult to ford. This is particularly true of the Rio Seco-Guay-jiniquil area, where there are no inhabitants and few open trails through the tropical jungle. Santa Cruz and Sardinal are served by the Taca Airlines, two planes a week from San Jose and Puntarenas. The port of Puntarenas has excellent loading facilities but none are available nearer the manganese districts; there ore must be lightered to vessels standing offshore. Physiography In Costa Rica there are three physiographic divisions: (I) the central plateau (Mesa Central), a somewhat dissected lava platform upon which are superimposed volcanic peaks, some of which are still active; (2) the Atlantic lowlands and (3) the Pacific lowlands. The peninsula of Nicoya and the larger part of the province of Guanacaste lie within the Pacific low- lands. The term "lowland" indicates relation to the central plateau, as the peninsula of Nicoya is largely mountainous. The mountains of Nicoya form the southern extremity of a range of low hills, which are almost continuous along the Costa Rican coast to the Nicaraguan border. The coast line shows evidence of subsidence. The Gulf of Nicoya obviously is the drowned valley of the Tempisque River. The west coast of Nicoya shows many rocks, shoals, bays, inlets and small islands as evidence of a sea encroachment. The altitude of the manganese deposits ranges from 0 to 800 ft. Topography is of the mature type of wide valleys and low hills, often with abrupt slopes. The climate is tropical and humid; rainfall of 125 in. falls largely between Sept. I and Jan. I. Vegetation often is very dense, the Rio Seco-Guayjiniquil area being truly tropical jungle, containing many hardwoods. History and Production The production of manganese ore in Costa Rica apparently dates from the first World War. It reached a peak of about 10,000 tons in the year 1919, then declined rapidly. Evidence of this period of prospecting has rotted away or been swallowed by the jungle. The war now in progress apparently revived the interest in Costa Rican manganese and has encouraged further prospecting. Deposits long known have been redenounced and trails have been cut through the jungle to them. Many new deposits have been discovered in the last

Jan 1, 1949

By A. E. J. Engel

INDUSTRIAL talcs and ground soapstones both include earth materials of different chemical and mineral compositions. In general, the industrial talcs are composed of silicates that contain appreciable amounts of magnesia. Generally talc, the mineral, is a predominant or at least common constituent. Talc, the mineral, is a hydrous magnesium silicate, with a suggested theoretical formula H2Mg3 (SiO3)4 or H2O.3MgO.4SiO2, which postulates 63.5 pct SiO2, 31.7 pct MgO, and 4.8 pct H20.11 Actually different single grains or leaves of pure talc have variations in water content ranging from 3.5 pct to more than 7 pct, as well as slight differences in the proportions of MgO and Si02.16 Only a very small percentage of the ground industrial talcs produced in the United States approaches the composition given above (see Table 1). The purer talcs invariably occur as alteration products of magnesium carbonate rocks. Some of these relatively pure talcs include compact, massive types largely mined in California, which are designated as steatite.38,58 In general, steatites cannot contain in excess of 1.5 pct CaO, 1.5 pct combined FeO and Fe203, and 4 pct A1203. The less pure varieties of massive talc and mixtures of magnesian silicates are commonly called soapstones, or merely talc. Many industrial earth products classified as talc or as soapstone contain a variety of minerals, especially serpentine (H4Mg,Si2O9), chlorites (hydrous magnesium, iron, aluminum silicates), tremolite (Ca2Mg5H2[SiO3]8), anthophyllite ([MgCa]SiO3), diopside (CaMg[SiO3]2), and quartz (SiO2). The carbonates calcite (CaC03), dolomite (CaMg[CO3]2), and magnesite (MgCO3) are also present in some industrial talcs and soapstones. Some of these minerals, as well as sulphides, sulphates, and iron and manganese oxides, if present at all or in excess of minute amounts, may alter the composition of the industrial talc sufficiently to restrict materially its use or to lower or destroy its value. These factors are discussed in a subsequent section of this chapter.

Jan 1, 1949

By Terkel Rosenqvist

Few subjects in iron metallurgy and steelmaking have caused so much discussion and have resulted in so many divergent and partly contradictory statements as the theory of the distribution of sulphur between molten iron and slag. An interesting survey of the different points of view is given by Grant and Chipman.1 Grant and Chipman themselves conclude in their paper that, at least in the open hearth and electric steel furnaces, the governing reaction is a simple distribution like: FeS steel ? FeS alas [Ia] and MnS steel ? MnS alas [Ib] They conclude that no significant reaction takes place according to the more classical scheme: FeS steel + CaO slag ? FeOslag + CaS slag [2a] This is the reaction which is regarded by most investigators as the primary sulphur reaction in the blast furnaces, where it is accompanied by the reaction with carbon: FeO slag + C ? Fe + CO [2b] The writer suggests an alternative to the reactions mentioned above. It is the opinion of the writer that the a priori assumption of defined molecular compounds being present in molten steel as well as in molten slag, has retarded rather than promoted a solution of the sulphur problem. As for the constituents of molten steel, it is taking too much for granted to mention the presence of FeS and MnS. If this infers that the melt contains diatomic molecules, the statement must be made that at least such molecules do not exist in the solid state at room temperature. Solid FeS crystallizes in the nickel-arsenide structure, and MnS (high temperature form) in the sodium-chloride structure. In these structures the sulphur is surrounded by six metal neighbors, and no molecular groups can be isolated. When sulphur is dissolved in molten steel, it is expected to maintain the same coordination number as in the solid state. We know that sulphur in steel has a certain chemical potential and a chemical activity which can be expressed by the reaction: S steel + H2 ? H2S where the activity A. is proportional to the ratio PH2S/Ph2. We know also that this activity is influenced by the presence of other constituents. For instance, there is reason to assume that carbon and silicon increase the sulphur activity and manganese decreases it. However, the exact relationship here is scarcely known, and further investigations are needed. Concerning the molten slags, if anything should be assumed a priori it would be that they are composed of ionic rather than molecular parts. There are indications of this in that solid silicates form ionic structures, and molten slags show high electric conductivity and can be electrolyzed.2 It has been claimed by Taylor and Chipman3 that their activity investigations

Jan 1, 1949

By H. C. Stumpf, L. A. Willey, V. L. Fink

Very few studies of the aluminum-sodium system have been reported. Heycock and Neville1 were unable to detect any solubility of sodium in liquid aluminum. Mathewson2 prepared an equilibrium diagram consisting of two horizontal lines at the melting points of the two metals. He was unable to detect solubility at either end of the system. Scheuer3 first developed a method for the chemical determination of sodium in aluminum and then studied the solubility of sodium in liquid aluminum. His method involved heating aluminum with excess sodium under hydrogen until the two liquids were at equilibrium, chilling the melt, and analyzing the aluminum-rich layer. He reported the solubility to be 0.10 pct sodium at 700°C, 0.115 pct at 750°C, and 0.128 pct at 800°C. This paper on the aluminum-sodium system is the twenty-first in a series from the Aluminum Research Laboratories describing equilibrium relations in aluminum alloy systems. Materials The alloys for this investigation were prepared from electrolytically refined aluminum (99.99 pct) and sodium of reagent quality conforming to specifications of the American Chemical Society. Chemical analysis of the aluminum showed the fol- lowing impurities: 0.002 pct Si, 0.002 pct Fe, 0.002 pct Cu, <0.001 pct Mn, 0.002 pct Mg and 0.001 pct Na. The sodium content of all alloys was determined by either chemical or spectro-graphic standard procedures5 No. IIA and No. SI I respectively. Determination of Miscibilitv The boundary of the miscibility gap was determined by a modification of the method of Scheuer3 and of that used by Henry and Cordiano. As finally developed the method was as follows: A fused alumina crucible 25 mm id with 75 mm depth and 1.5 mm wall thickness was placed on the bottom of a container made from a 4-in. length of 1 1/2-in. iron pipe and pipe caps. ?&apos;he space around the sides of the crucible was filled with 40 mesh crystalline alun-dum. A charge of 60 g of aluminum and 10 to 15 g of sodium was placed in the crucible, the sodium first being trimmed and rinsed with iso-pentane to remove kerosene. The container was tightly capped, filled with argon (99.6 pct pure), and heated to the desired temperature (between 670 and 800°C). A positive pressure of argon (about 8 mm of mercury) was maintained through a small pipe to the upper cap. After soaking for from 2 to 5 hr, the container was quenched by directing a stream of water onto the base, so that solidification of the aluminum-rich layer was rapid and essentially unidirectional. Analyses were made on drillings taken from the bottom of the ingots after machining off the surface layers. Pre-

Jan 1, 1949

By H. L. Talbot, R. S. Young, A. Golledge

The differentiation of oxidized forms of lead from lead sulphide in complex products by chemical analysis is of considerable importance to certain mining and metallurgical companies. A method for the separation of "oxide" lead, (non-sulphide), from sulphide lead, based on the solubility of the former and virtual insolubility of the latter in concentrated ammonium acetate solution, has been employed for many years in the industry.1,2 Unfortunately this solvent, while satisfactory for lead sulphate and carbonate, is without action on lead phosphates or vanadates, with the result that at .Rhodesia Broken Hill Development Co. where these latter minerals are present, they have counted as sulphides when using this procedure. For the control of operations in the new concentrator it became necessary to have a closer check on sulphide lead than could be obtained by the only current method of differentiating this constituent from oxidized lead. For some time this problem has engaged our attention, and after a large number of solvents had been tried, the following procedure was evolved. Like all methods depending on selective solution in a com- plex material, it will not give 100 pct separation into oxide and sulphide portions. Because of the ease of oxidation of galena this is almost an impossibility. Our procedure will, however, give a recovery of 99-100 pct of the oxidized lead, contaminated with very small quantities of lead sulphide, and a corresponding recovery of 97-98 pct of the sulphide lead. The procedure is simple and reproducible, and gives results quite .satisfactory for routine control purposes. Experimental For this work standard samples of the following pure lead minerals were obtained from Broken Hill, Northern Rhodesia. These were hand-picked specimens, and were carefully ground to —200 mesh. Since it was realized that a fair-sized sample would be necessary for this work to thoroughly test the large number of procedures we had in mind, a compromise had to be effected between a few grams of very pure minerals and a large sample of lower grade material. For this work the presence of gangue was not deleterious, so long as the oxide lead minerals did not contain lead sulphide, and vice versa. Table I shows the lead minerals used, and their total lead content, both theoretical and actual. Apart from the pyro-morphite the minerals are seen by analysis to be quite pure. Many solutions were tried in an effort to dissolve the oxidized lead minerals and

Jan 1, 1949

By H. P. Hermance

Preliminary examinations of the Parker and Webb deposits were made by E. E. Maillott, Bureau of Mines engineer in July 1942. The chemical and spectrographic analyses included in this report wore made at the Eastern Experiment Station of the Bureau of Mines at College Park, Md., under the general supervision successively of S. M. Shelton and of Paul M. Tyler, formerly regional engineer, Eastern Region. Beneficiation tests were conducted under the supervision of Frank D. Lamb. The mineral exploration was under the general supervision of J. D. Bardill, formerly district engineer. Acknowledgment is gratefully made to the St. Joseph Lead Co., in particular to John S. Brown and P. J. Mechin for their assistance. Special acknowledgment is made to the American Metal Co., Ltd., ,Ind to T. C. Moore for furnishing core-drill and assay data; to the Universal Exploration Co. and P. B. Jennings; and to the Northern New York Mining Corp. and C. H. Sprague for their permission to publish the core-drill dta. Excellent cooperation was given by Dr. W; M. Agar and C, L. Rogers of the Geological Survey.

Jan 1, 1949

By C. O. Swanson

Introduction It is thirty-odd years since I first entered the mining field by working underground at Anyox, which makes 1t clear that I can qualify as one of the older generation and claim the privilege of dealing in generalities and broad impressions. In this position I find it interesting to re~ view the salient changes that have taken place during the past quarter-century and indulge in some speculations concerning the future, with the hope that my comments may be of value to others engaged in the field of mining geology. As indicated by the title, the treatment is a perspective from a single viewpoint rather than an attempt to give a complete and impartial analysis. This plan, while deficient in some respects, has the advantage of being more specific than one based on .a. compilation of opinions. An additional advantage is brevity, which would be difficult to obtain on a subject of this scope if it were viewed from all angles. In conformity with the general plan, specific references to the literature have been avoided as much as possible, although a number of publications of general nature are listed at the end of the paper. This procedure is intended to give credit to the sources used without making them in any way responsible for the interpretation. The starting point for the discussion is the situation as it existed twenty-five years ago, when, as a recent graduate, I first had a fairly complete picture of it. However there is one impression of an earlier; stage that seems worth mentioning. In one of my first courses in geology the statement was made that some of the leading authorities of the previous generation, who guided geological thought at about the beginning of the century, were inclined to question the value of an association with mining. This attitude was no doubt exceptional, and certainly it was not adopted by the students who laid the groundwork for mining geology in North America. But its existence even in a limited circle is significant m showing how far the application of geology has advanced as the result of the investigations started during the first part of this century.

Jan 1, 1949

Jan 1, 1949

By Enid C. Plante

To extend our knowledge of the flotation behavior of sulphide minerals, the response of the following minerals to ethyl xanthate as collector was studied by captive bubble and cylinder flotation tests: antimonite (Sb2S3), arsenopyrite (FeAsS), covellite (CuS), rnarcasite (FeS2), orpiment (As&.), pyrrhotite (FeS) and tetrahedrite (4CU2S, Sb2S3). Lollingite (FeAs2) was also exarnined to determine if its behavior could be correlated with that of pyrite or rnarcasite (FeS2. Experimental Method Mineral for Flolation and Contact Tests Mineral was prepared in the manner a1ready described1 and its behavior both freshly ground and "oxidized" was studied. Tests in which the mineral was allowed to oxidize for various periods showed that a uniform response to xanthate was reached in two days or less. Oxidation Products The soluble oxidation products from most of the above minerals were measured and their effect on the flotation of the mineral studied. The mineral samples were wet ground to minus-52 mesh and deslimed, and 20 g of this material placed in 200 ml of distilled water of a measured pH value, and CO2-free air bubbled through each suspension at a constant rate (about I bubble/sec) for 5 days. It has been shown1 that the size of the mineral is unimportant in the rate of formation , oxidation products and that probably the supply of oxygen to the mineral is the rate-determining step. At the end of 5 days the pH value of the &apos;supernatant liquid was measured again, the mineral filtered off and the soluble products determined. Methods for estimation of sulphur acids, ferrous and ferric iron and copper have already been described.l No satisfactory method was found for antimony. Arsenic was estimated as the arsenate.2 Flotation of "Unoxidized" Minerals with Ethyl Xanthate as Collector Fig I to 3, 4. 6, show the effect of alkali and cyanide as depressants for the minerals with 25 mg potassium ethyl xanthate per liter as collector at 35°C. Wark and Cox8 have shown that with ethyl xanthate as collector, galena, pyrite and chalcopyrite show a greater tolerance to cyanide and hydroxyl ions at room temperature than at 35°C Covellite (Fig I), marcasite (Fig 2) and pyrrhotite (Fig 3) behave similarly. The independence of the curves for pyrrhotite at 35°C with respect to cyanide

Jan 1, 1949

By D. M. Davidson

The drill sampling and evaluation of the Piney River ilmenite property was carried out during the early part of 1944 by E. J. Longyear Co., using the company&apos;s standard core-drill outfits to complete 44 holes to an average depth of IIO ft., from which were collected 1100 samples. The Piney River ilmenite-apatite lode is 25 miles north of Lynchburg, in Nelson and Amherst Counties, Virginia (Fig. I). Its elevation is about 900 ft. above sea level and it lies in the old pre-Cambrian complex of crystalline rocks that make up the Piedmont physiographic province lying east of the Blue Ridge Mountains. The terrain is gently rolling and well watered and is drained by the Piney and Tye Rivers, which are tributary to the James. Geology The Piney River ore body comprises the upper, weathered part of a tabular basic rock, which occurs in a large area of granite-gneiss. The section containing profitable ore is about 3000 ft. long. It is separated naturally by thinning, into northeastern and southwestern segments with widths of 300 and 100 ft. respectively. The dip averages 4.5° southeast but steepens locally (Figs. 2 and 3). The unaltered rock is an even, granular mixture of feldspar, ilmenite, and apatite, to which Watson1 gave the name Nelson-ite. The decomposed ore body, which is now being exploited by open cut, is a yellowish clay material in which black ilmenite and glassy apatite are the only conspicuous minerals. An economically important structural feature of the deposit is a banding of minerals coinciding with the dip and strike of the dike. This is more pronounced in the footwall of the northeastern part of the intrusive, where there is also a higher than average concentration of ilmenite. According to Watson,&apos; the TiO2 content of the Nelsonites of this region ranges from 10 to 42 per cent, while P205 varies from 4 to 15 per cent. It is inexpedient at present to disclose exact metal content of the Piney River ore body; suffice it to say that the average tenor lies well within the extremes noted by Watson. Sampling Methods and Procedure During a preliminary inspection of the ore body, Longyear engineers surmised that because of its soft, clayey character, drill sampling would yield little core, therefore they made all possible preparations to procure accurate samples of the drill cuttings, or sludge. The first requisite in obtaining representative drill cuttings is to secure their full recovery. Only under certain conditions, and by employing a most careful and rigid sampling technique, is this pos sible. There must be no loss of the water circulating in the hole, and the velocity

Jan 1, 1949

By Raymond E. Birch, Oscar M. Wicken

THE mineral magnesite, formerly the source of nearly all magnesia, now shares this role with brucite, dolomite, and the world&apos;s natural and artificial brines. The mineral magnesite is the normal carbonate of magnesium, MgCO3, composed when pure of 47.6 pct MgO and 52.4 pct CO2. It is one of the calcite (CaCO3) group of rhombohedra1 car- bonates, which also includes dolomite, CaCO3.MgCO3, and siderite, FeCO3. Magnesite does not form an isomorphous series with calcite or dolomite; consequently, these lime-bearing carbonates commonly present in magnesite deposits are there only as mechanical mixtures. Magnesite and iron carbonates form a continuous isomorphous series and therefore are inseparable mechanically. The variety of magnesite known as breunnerite is an isomorphous mixture of iron and magnesium carbonates. Magnesite may be either crystalline or amorphous (cryptocrystalline). The former, which includes breunnerite, is often termed spathic magnesite. Crystalline magnesite varies from finely to coarsely crystal- line in texture and has a hardness of 3.5 to 4. The color is white, yellowish, blue-gray to drab, red, pink, black, or mottled. It is seldom found pure but usually contains variable amounts of iron, lime, silica, and sometimes manganese. The color cannot be used as an index of purity. The cryptocrystalline variety is perhaps the more common type but the crystalline variety usually occurs in larger deposits. The amorphous type is fine grained and compact, showing no cleavage. It is usually snow white but sometimes is light to pale orange yellow or buff, owing to impurities. The fracture is conchoidal to uneven and the hardness 3.5 to 5. Silica may be present as inclusions of serpentine, quartz, chalcedony, or other minerals. The lime and iron contents are usually low. In general, amorphous magnesite is somewhat purer than the crystalline variety. The specific gravity of cryptocrystalline magnesite is 2.90 to 3.00. Pure crystalline magnesite has shown a specific gravity of 3.02 but iron

Jan 1, 1949