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By J. R. Cahoon, W. V. Youdelis

In a recent publication by Flemings and Nereo, equations for the calculation of the solute distribution in castings were developed, and the solute distributions for unidirectionally solidified A1-4.5 pct Cu ingots cast under various thermal conditions were calculated. The equations derived by Flemings and Nereo for the dendrite "expansion" and "steady-state" zones are essentially equivalent to those previously published by Kirkaldy and Youdelis, 16 for, although the flow velocity cooling rate, and temperature gradients appear in the fundamental differential equations, calculation of the solute distribution requires that the above be explicitly related to the assumed or measured solid-liquid mass distributions during solidification. In the treatment of Kirkaldy and Y-oudelis the mass distributions are simply assumed or measured without any prior and explicit association of the mass distributions to a velocity parameter. The treatment of Flemings and Nereo permits solidification conditions more complicated than the simple unidirectional case to be considered (although in principle the treatment of Kirkaldy and Youdelis could also be extended to include the more general solidification case); however, we note some limitations in the treatment of Flemings and Nereo, and the purpose of this discussion is to point these out and also to indicate those points of similarity in the mass distribution models used which make the two treatments equivalent. Both the Kirkaldy-Youdelis and Flemings-Nereo treatments consider a zone of dendrite "expansion" followed by a "steady-state" zone. However, Flemings and Nereo have failed to consider a zone of "collapse" which begins at the base of the first dendrites to reach the top of the ingot. The omission of this zone of "collapse" results in too high a value for the solute concentration in the latter portion of the ingot to solidify and the conservation of total solute is not possible. In the zone of "collapse" the solute concentration rapidly decreases, which may be accounted for as follows: consider a mushy zone of length X,, shown schematically in Fig. 11 (a), which develops at the chill face and then moves as a steady-state configuration along the length of a solidifying ingot. The solute distribution in this zone, when it first attains the steady state, is shown qualitatively in Fig. 11(b). The segregation is positive (above Co) in the region ab because this region contains primary dendritic solid, ranging from kco to kcE in composition (where k is the partition coefficient and CE is the composition of the eutectic liquid), and liquid metal near eutectic composition. The segregation is negative (below Co) in the region bc because it contains primary dendritic solid only slightly above kco in average composition and liquid metal only slightly above Co in composition. As the zone X, moves as a steady state

Jan 1, 1969

By Franklin Moeller

CONSIDERING the really short time that has elapsed since hydro-metallurgical processes of extracting copper from ores have been extensively developed, and the large scale on which this method is practised at Ajo, the successful operation of the plant must excite the admiration of every visitor. The application of the mechanical unloader to the purpose for which it is employed at Ajo is but an extension of the use for which it was originally designed, some 20 years ago, for unloading iron-ore vessels at the lower lake ports. At a time when the maximum output of a single machine, with men shoveling the ore into buckets, did not exceed 40 tons per hour, G. H. Hulett invented and constructed a mechanical unloader with a bucket of 10-ton capacity. This machine was of the steam hydraulic type, and it is still in operation after nearly 20 years of service. Several more machines of the same type were built, but the difficulties of power transmission, and the rapid development of electrical machinery, soon led to the adoption of electrically operated machines, and practically all of the later installations have been thus equipped. This electrically operated unloader has proved so well fitted for the work that the design of the lake boats has been practically revolutionized, with special view to the ease and rapidity of unloading. It is now possible to unload a 12,000-ton vessel with four machines in a little over 2 hr. This has necessitated an increase in the capacity of the machine, and the largest yet built handles an average load of 17 tons, with a possible maximum of 20 to 21 tons.

Jan 8, 1918

By F. Wartenweiler

Much has been written on milling and cyanidation as practiced on this gold field, therefore the purpose of this paper is to record its development only briefly and to describe in more detail the contemporary state of the science and art of this branch of metallurgy. It is a pleasing task, because a steady evolution is to be recorded. Definite trends influenced the development. Some have the stamp of original thought and observation; many have been achieved from a logical sequence of planned experimentation and research; while others have resulted from the drive of sheer necessity. The large scale of operations and capital outlay has compelled conservative progress. Banket ORe For a better understanding, a review of the characteristics of the ore may not be out of place. The Witwatersrand gold ore is named "banket" from the similarity of the oxidized outcrop ore to the Boer almond cake of that name. The inter bedded orebody is commonly called "reef." It is a conglomerate composed of various sizes of pebbles, principally non-gold-bearing white quartz, which constitute about 70 per cent of the bulk, within a matrix of cementing material consisting of fine-grained quartz, sericite, chlorite and chloritoid and containing the precious metals and mineral sulfides. Argillaceous components in the mil pulp may be as high as 32 per cent. Of the sulfide minerals, pyrite and pyrrhotite are predominant. Most base-metal sulfides are present, but sporadically and generally only found in highly concentrated mill products. In the latter, osmiridium is recovered in varying amounts; from 1 oz. in 3000 tons to 1 oz. in 10,000 tons of ore. Most of the oxidized ore has been mined, so milling is confined principally to the sulfide ore. A typical analysis of sulfide ore is: silica (SiO2), 86.76 per cent; pyritc (FeS2), 2.75; ferric oxide (Fe2o3), 2.65; alumina (Al2o3), 6.91; lime (CaO), tr.; magnesia (MgO), 0.70; sp. gr., 2.7. Milling grade at present price of gold (137 shillings per ounce) varies from 3.5 to 8.5 dwt., with one mine milling ore as high as 16 dwt. The

Jan 1, 1935

ARSENIC and orpiment are two mineral substances of a similar kind, and they are by nature pure, without admixture of other species. In regard to their apparent qualities we shall say that their composition is scorched, well-purified earth; and because of a certain subtlety and easy digestion they penetrate very easily into fused metals, indeed they act in such a way that they corrupt and convert almost into another nature any metal with which they find themselves. Orpiment and arsenic act in almost the same way as do tin and mercury. With these the fraudulent alchemists blanch copper, brass, and even lead to the whiteness of silver. According to the physicians they are in fourth part hot and dry. They are also corrosive because of certain powers; indeed, they are a most powerful poison to the life of all things. Arsenic is found in white and citrine colors. Orpiment seems to be a separate thing, although both are formed in the same ore and both are in lamellae or scales. These lamellae or folds are thinner than paper, in fact

Jan 1, 1942

Located 19 km (12 miles) north of the Carlin gold mine and 43 km (27 miles) north of Carlin, Nevada, is Carlin Gold Mining Company's Bootstrap mine. This deposit is similar in character to the Carlin ore, except it occurs in the upper plate formation of the Roberts thrust fault rather than in the lower plate rocks. The deposit was discovered over 50 years ago and operated on a small scale in the early 1960s using vat cyanide leaching. Newmont Mining Corporation acquired the property in 1967, and from an open pit mined 650,000 mt (717,000 st) of ore which was trucked to the Carlin conventional cyanide plant. (See Gold and Silver Cyanidation Plant Practice by F. W. McQuiston, Jr. and R. S. Shoemaker, Society of Mining Engineers, AIME, 1975. ) With the same grade of ore, the gold recovery by conventional methods is somewhat lower than that of the Carlin ore, even with a finer grind. Leaching tests were first conducted by the United States Bureau of Mines in 1968; and in 1971, ‘72, Carlin conducted extensive cyanidation investigations by bottle testing, simulated heap leaching in tall columns, and small scale pad heap leaches. Gold extractions ranged from 55% to 95% with low reagent consumptions. In 1974, a 545 mt (600 st) heap leach test was made at Carlin on pit-run low-grade material assaying 1.37 g per mt (0.04 oz per st). In a six week period the extraction of gold was 47.57'0. Carlin uses atomic absorption analyses on cyanide leach liquors to determine potential leach ore grade rather than fire assaying because it gives a more practical assessment of leach recovery than does fire assaying, which accounts for the total gold content. Before open-pit mining was started, plans were made to heap leach, the ore at Bootstrap which was too low grade for transport to Carlin. The topsoil outside of the designed pit area overlaid ancient lakebed clays which were found suitable for compaction to form an impervious base on which the dumps to be leached could be placed. By using the cost of truck haulage and milling at the Carlin plant and the price of gold 'at that time, a cutoff grade was established which resulted in the gold content of the first 150,000 mt (165,000 st) of the heap leaching ore being 1.51 g per mt (0.044 oz per st). An additional 500,000 mt (552,000 st) of lower grade material, assaying 0.96 g per mt (0.028 oz per st), is available for leaching as economics permit.

Jan 1, 1981

By R. S. Sherwin

The extractive metallurgy of primary aluminum from its ores is discussed with special attention to the production of alumina from high grade ores by the Bayer process, including differences between American and European practice and a brief description of some processes for lower grade ores and the electrolytic reduction of the oxide to aluminum. METALLIC aluminum is not found in nature, but the oxides, hydroxides, and especially the silicates are plentiful. The estimated percentage of in in the crust of the earth is about 8 pct while that of iron is about 5 pct. By far the larger portion of this is combined with silica in the form of various clay minerals and igneous silicate rocks. From the point of view of extractive metallurgy of aluminum, these are low grade ores while the better qualities of bauxite are the high grade ores. There have been various definitions of bauxite but perhaps the best definition at the present time is that bauxite is a rock or earth commonly used as an ore of aluminum or its salts in which the aluminum is present predominantly as a hydrate or a mixture of hydrates and hydrous oxides. It contains varying amounts of oxides of silicon, iron, and titanium and traces of compounds of some of the less common elements. The silica is mainly combined with alumina as clay or clay minerals which are hydrous aluminum silicates, although a part of it may be present as quartz sand. On the American continents, the alumina is mainly present as gibbsite, Al2O3 . 3H2O, and the same may be said of the best known deposits of the Dutch East Indies and some of the deposits in India. In France and other countries in Europe as well as in Africa, the alumina is present mainly as boeh-mite, A12O3 . H2O, but in some of these deposits it is mixed with minor amounts of gibbsite. Some other deposits, such as those in the islands of Haiti and Jamaica, evidently contain two or more hydrates or hydrous oxides of alumina in varying proportions. Perhaps the main portion of the alumina may be present as gibbsite and boehmite with the proportion between the two varying rather widely. In the silicate minerals, including clay, the alumina is chemically combined with silica and has not been separated satisfactorily by mechanical or physical ore-dressing methods. Low grade bauxites are mixtures of hydrates, usually gibbsite or boehmite, with clay, iron oxides, etc. In some low grade bauxites, it is possible to separate a portion of the gibbsite or boehmite, which may be present as relatively hard and coarse particles, from soft or finely divided clay minerals by log washing or similar methods. This has been applied to the product of some mines or parts of them, but on other ores it is not applicable. In some cases the gibbsite or boehmite is almost as fine and soft and of nearly the same specific gravity as the clay minerals so that washing and gravity separations are not successful. The iron oxide, the clay minerals, and a part of the titanium minerals are often so finely dispersed in the ore that any of the physical mineral separation methods, including separations by gravity, particle size, flotation, and electrostatic or magnetic separation, have not been commercially SUCCESSFUL except on relatively small lots of ore. For these reasons, the only available methods of separation on the general run of ores have been methods which would be classed as chemical rather than physical or mechanical separations. Aluminum oxide can be reduced by carbon at temperatures above 1800°C to form metallic aluminum and aluminum carbide or nitride. The temperature for rapid reduction of aluminum oxide to metallic aluminum is about the boiling point of aluminum and above the temperatures necessary to reduce iron, silicon, and titanium so that the direct reduction of an aluminum ore with carbon will produce an alloy of aluminum, iron, titanium, silicon, etc., which may be mixed with carbides and nitrides. Also a large amount of the reduced aluminum may be lost as a vapor except in the presence of some alloying agent such as copper or other metals. While it is possible to refine such alloys or mixtures so as to produce commercially pure aluminum, the methods which have been found are too expensive for the present market. One direct reduction method which found limited commercial use in Germany during World War II was the direct reduction of a mixture of clay containing very little

Jan 1, 1951

THE BULOLO gold-dredging operations were described at some length by Frank W. Griffin at the Oct. 11 meeting of the San Francisco Section. This enterprise is on the island of New Guinea, about 1800 miles north of Sydney, Australia. The area in which the present placers exist originally was German territory but later the location was taken over by Australians. Gold was discovered in 1923 by prospectors who required 8 to 9 days to pack into the country over severe mountain trails. In 1926 a rich discovery was made on Edie Creek and in 1927 the deposit on the Bulolo River was discovered.

Jan 1, 1932

By -1ng. B. Stampa

Introduction The author investigates the technology available for mine flow reversibility by reversing the operation of the main fans. Air bypassing, reversed sense of impeller rotation, and blade angle overturning are indentified as the most important means of achieving the flow reversibility. The axial flow fan is studied in some detail and the important variables of fan capacity, pressure, and efficiency are compared. Although some references to rotor velocity vectors are made (Fig. 1) the subject is not presented in detail in order to stay within the paper's intent. Background Where applicable, the requirements of mine air reversibility are an important part of the overall mining operation and should be given proper attention during mine ventilation design. The need for preventing mine fires in intake shafts is the reason that, for a long time, it has been a requirement in the Eastern European countries to have the feature of flow reversibility built into all main mine ventilators. In West German coal mines, this requirement does not apply since higher priority is given to safety precautions. Some metal and mineral mines in the United States are required to have ventilation currents readily reversible in order to fight underground fires as stipulated by the Title 30, Mineral Resources, and by the mining laws of some states. From an equipment performance standpoint, assuming the validity of other flow technical data, the question of the ventilator flow reversibility is of ten related to the actual scenario of the begining of an emergency. Most of the time, the system operates within half of the design capacity only and assuming that the equivalent mine orifice is stable only a quarter of the design pressure must be produced by the fan in any event. For such an operating point, the effective gradient and energy losses can be disregarded since fire emergencies usually last for only a short period of time. Equipment Comparison Fig. 2 depicts a ventilation arrangement wherein the flow can be reversed through changing the functional position of the dampers (slideplate or blade type). Both axial or radial fans can be applied in this case. For the mine exhaust operation, slide plates (guillotines) I and I11 are opened while plates I1 and IV are closed. The flow thus encounters two reroutings, which due to the reverse ventilation flow cannot be equipped with proper sheet metal deflectors, and therefore large pressure losses result during the air reroute operation. When the intake operation is considered, the dampers I and I11 are closed and the pair I1 and IV is opened. For this arrangement, almost the same flow conditions exist as for the foregoing exhaust mode; however, one additional corner must be applied. Even if a highly efficient ventilator is used in one of the previously described systems, the additional pressure losses resulting from air rerouting plus those encountered between the planes "0" and "00" significantly affect the overall energy requirements. Illustrated in Table 1 are the relative component losses of even flows for evaluation of various detour arrangements. From the aerodynamics standpoint, when these components are being designed, serious consideration must be given to the individual levels of the pressure loss coefficient, 5. Should there be an increase in the sum of the loss coefficients due to the particular component selection such that, for example, the net change is 1.2; the resulting pressure to be produced by the fan can be calculated from:

Jan 1, 1981

By D. E. Peacock, B. Harris

The mechanical behavior of a niobium (columbium)TUNG alloy containing 20 wt pet Ta. 15 wt pet W, and 5 wt pct Mo has been studied in the temperature range 77° to 423°K. All specitrzens tested, apart from those into which oxygen was introduced deliberately, were fully recrystallized at 1500°C, had an average grain diameter of approximately 0.03 mm, and a combined carbon, ox-vgen. nitrogen, and hydrogen content of not not more than 250 ppm The meckatnical behavior of this alloy was found to be similar in seine respects to that of unalloyed bcc transilion metals. Both yield and flow stresses were sensililae to changes in deformation rate and both increased rapidly as the temperature was lowered below 150°C. A detailed examination revealed, however, that most of the temperature dependencc of the yield stress could be accounted for by the temperature dependence of the elaslic, modulus, whish is not true of the unualloyed metals. In lension at moderate strain rates, the trunsition fvom ducfile to hvittle behavior occurred just belour room teunperature bur incveasing the oxygen content from 190 to 200 pptn raised the transition temperature to about 4000K. TUNGSTEN and molybdenum have been found to be very effective high-temperature solid-solution strengtheners of niobium, but the amounts of these elements that can be added are limited if the alloy is to retain useful room-temperature ductility. It has been shown, however, that partial replacement of niobium with tantalum lowers the ductile-to-brittle transition temperature of such alloys and allows larger additions of molybdenum and tungsten to be made. Some of the high-temperature properties of niobium alloys containing these elements have been reported by Bartlett et al.,' and the work presented here describes the low-temperature behavior of one of these, a solid-solution alloy containing, by weight, 20 pct Ta, 15 pct W, and 5 pct Mo. The elastic modulus, the temperature and strain-rate dependence of the yield and ultimate tensile stresses, and strain-hardening effects are discussed and compared with those of unalloyed niobium. EXPERIMENTAL MATERIAL AND PROCEDURE The material used in this investigation was produced from a 6-in.-diameter, double arc-melted ingot. This was can-extruded at about 1150°C to a 3-in.-diameter billet, and after annealing for 1 hr at 1345°C the billet was heated to 1315OC and part rolled and part forged to 0.5-in.-thick plate. The resulting material exhibited a cold-worked structure

Jan 1, 1965

By T. W. Smoot

Bentonites are composed chiefly of montmorillonites which are clay minerals generally distinguished from other clay minerals by their surface activity and extremely fine particle sizes. The uses of bentonite are many and varied, depending principally on one or more of the following characteristics: 1) Thixotropy, 2) Dispersion, 3) Plasticity, 4) Swelling, 5) Surface area, activity, ab- and adsorbtive properties. Some of the major uses of bentonite are briefly reviewed as are the mineralogical and physical properties. The uses of X-ray diffraction, differential thermal analysis and other tests for physical properties are reviewed as to their uses in evaluation. In the past 40 years, the industrial uses of bentonite have grown to include many applications. The physical and chemical properties of bentonites are such that they can be used as catalysts, adsorbants, plasticizers, binders, drilling muds, and pelletizers. Not all bentonites, however, can be used for all of these purposes and some are probably of little or no value for any use. The purpose of this paper is to briefly review the mineralogical, physical, and chemical characteristics and variations of bentonites. The determination of these characteristics and their uses in industry are also reviewed so that these determinations can be used as guides for evaluation. DEFINITION OF BENTONITE The word bentonite, as used in this paper, refers to a clay material, usually (but not necessarily) the result of the in-place-weathering of volcanic ash. Its mineralogical composition is dominated by the clay mineral montmorillonite and generally carries other clay minerals and silica minerals as the major 'impurities'. For this discussion, such a definition is justified since most of the economic uses of bentonite~ depend upon the physical and chemical characteristics of the montmorillonite minerals involved. The terms montmorillonite and bentonite have been used almost interchangeably in industry although the latter is much more commonly used. Therefore, where montmorillonite is discussed, it should be understood that the major feature of a bentonite is the subject. STRUCTIJRE OF MONTMORILLONlTE The economic importance of bentonites is generally based on one or more of the following characteristics of montmorillonite: 1) high-thixotropy and easy dispersion in water, 2) high-plasticity, 3) swelling properties, 4) large surface area/unit mass, 5) high surface activity, and 6) adsorptive properties. To understand these properties, an appreciation of the structural scheme of montmorillonite is helpful. Montmorillonite IS one of the three-layer clay minerals composed essentially of hydrous aluminum silicates. The essential building blocks of these layered silicate:; are: 1) silica tetrahedron composed of a silicon atom surrounded by four oxygen atoms, and 2) aluminun (iron, magnesium, etc.) octahedral unit that, ideally, has one of these atoms surrounded by six hydroxal units. Fig. 1 is a diagrammatic sketch of the montmorillonite structure. In the montmorillonite structure, the tetrahedron are pointed inward and attached to the octahedral units so that the O= atoms replace the hydroxal units where they are bonded; hence two-thirds of the hydroxal units

Jan 1, 1962

By Richard W. Smith

There are three distinct varieties of phosphate rock, in Tennessee, known commercially as: (a) the "brown" rock, which is the residual product of the weathering and natural concentration of certain phosphatic Ordovician limestones; (b) the "blue" rock, which is an unaltered phosphatic stratum of Mississippian or Devonian age; and (c) the "white" rock, which is the result of chemical replacement and deposition. Before 1894, South Carolina and Florida had supplied most of the phosphate of the United States, South Carolina leading until that date and Florida from then until the present time. In December, 1893, the blue phosphate of Tennessee was discovered almost simultaneously on Swan Creek, in Lewis County, by two independent parties—by one as the result of a careful and systematic search for a workable deposit of the long known kidney phosphate, and by the other through mistaking it for a "bloom" of coal. Leases were taken and, with further prospecting disclosing other deposits in the region, the development progressed as rapidly as could be expected without railroad transportation. In January, 1896, Judge S. Q. Weatherly, who was interested in the development of the blue phosphate, while riding along the road south of Mt. Pleasant, noticed some slabs of a brown rock in place above the limestone. He took samples, and had them analyzed, thus discovering the high-grade brown phosphate of this district. This discovery was kept a secret until July, 1896, when mining was started at the village of Mt. Pleasant. Development proceeded rapidly, many owners obtaining their capital by the sale of their rock fences, which were largely composed of slabs of high-grade brown phosphate. The development of the brown rock, which is richer and more easily mined than the blue, caused Tennessee to advance from a poor third in production to second, surpassing the output of South Carolina, and being excelled only by Florida.

Jan 1, 1927

By Paul Tyler

STRATEGICALLY situated in almost the heart of the leading fertilizer-consuming area of the United States, Tennessee long has ranked second only to Florida as a phosphate-producing state. Since 1932 its share of the national output of phosphate rock has increased steadily, and in 1937 its production mounted to an all-time record. More significant even than the increase in tonnage are the forward strides in technology. The Tennessee Valley Authority's research program has led to quasicommer-cial operations using hydroelectric power to manufacture concentrated phosphatic fertilizer that is used to further the farmer education program designed to streamline the fertilizer business and halt the appalling waste of farm land by soil erosion. Even more stimulating to chemically minded imaginations is the production of elemental phosphorus by a private company that not only has succeeded in extracting it from the phosphate rock in electric furnaces but actually has contrived to capture it and to transport it safely in tank cars. New values have been created for low-grade phosphate deposits, and the economic life of reserves has been lengthened while the waste of former years goes back into the mills for recovery of its values by modernized methods. Tennesse's phosphate-mining industry dates from the discovery of blue rock in Lewis County in December 1893. Its known reserves are of the order of 101,500,000 tons, which compares with the production of 825,099 tons in 1937; that is, 21 per cent of the phosphate-rock output of the whole nation. The total output reported from 1894 until July 1, 1938, was approximately 19,000,000 tons. Official estimates of reserves are overly conservative and evidence submitted before a joint Con-gressional committee in November 1938 places "probable" reserves at more than 6,000,000,000 tons.

Jan 1, 1939

By Herbert R. Mosley, Paul M. Tyler

STRATEGICALLY situated in almost the heart of the leading fertilizer- consuming area of the United States, Tennessee long has ranked second only to Florida as a phosphate-producing state. Since 1932 its share of the national output of phosphate rock has increased steadily, and in 1937 its production mounted to an all-time record. More significant even than the increase in tonnage are the forward strides in technology. The Tennessee Valley Authority's research program has led to quasicommercial operations using hydroelectric power to manufacture concentrated phosphatic fertilizer that is used to further the farmer education program designed to streamline the fertilizer business and halt the appalling waste of farm land by soil erosion. Even more stimulating to chemically minded imaginations is the production of elemental phosphorus by a private company that not only has succeeded in extracting it from the phosphate rock in electric furnaces but actually has contrived to capture it and to transport it safely in tank cars. New values have been created for low-grade phosphate deposits, and the economic life of reserves has been lengthened while the waste of former years goes back into the mills for recovery of its values by modernized methods. Tennessee's phosphate-mining industry dates from the discovery of blue rock in-lewis County in December 1893. Its known reserves are of the order of 101,500,000 tons, which compares with the production of 825,099 tons in 1937; that is, 21 per cent of the phosphate-rock output of the whole nation. The total output reported from 1894 until July 1, 1938, was approximately 19,000,000 tons. Official estimates of reserves are overly conservative and evidence submitted before a joint Congressional committee in November 1938 places "probable" reserves at more than 6,000,000,000 tons.

Jan 1, 1939

By Paul M. Tyler, Herbert R. Mosley

Strategically situated in almost the heart of the leading fertilizer-consuming area of the United States, Tennessee long has ranked second only to Florida as a phosphate-producing state. Since 1932 its share of the national output of phosphate rock has increased steadily, and in 1937 its production mounted to an all-time record. More significant even than the increase in tonnage are the forward strides in technology. The Tennessee Valley Authority's research program has led to quasicommer-cia1 operations using hydroelectric power to manufacture concentrated phosphatic fertilizer that is used to further the farmer education program designed to streamline the fertilizer business and halt the appalling waste of farm land by soil erosion. Even more stimulating to chemically minded imaginations is the production of elemental phosphorus by a private company that not only has succeeded in extracting it from the phosphate rock in electric furnaces but actually has contrived to capture it and to transport it safely in tank cars. New values have been created for low-grade phosphate deposits, and the economic life of reserves has been lengthened while the waste of former years goes back into the mills for recovery of its values by modernized methods. Tennessee's phosphate-mining industry dates from the discovery of blue rock in Lewis County in December 1893. Its known reserves are of the order of 101,500,000 tons, which compares with the production of 825,099 tons in 1937; that is, 21 per cent of the phosphate-rock output of the whole nation. The total output reported from 1894 until July 1. 1938, was approximately 19,000,000 tons. Offcial estimates of reserves are overly conservative and evidence submitted before a joint Congressional committee in November 1938 places "probable" reserves at more than 6,000,000,000 tons.

Jan 1, 1942

By Herbert R. Mosley, Paul M. Tyler

Strategically situated in almost the heart of the leading fertilizer-consuming area of the United States, Tennessee long has ranked second only to Florida as a phosphate-producing state. Since 1932 its share of the national output of phosphate rock has increased steadily, and in 1937 its production mounted to an all-time record. More significant even than the increase in tonnage are the forward strides in technology. The Tennessee Valley Authority's research program has led to quasicommer-cia1 operations using hydroelectric power to manufacture concentrated phosphatic fertilizer that is used to further the farmer education program designed to streamline the fertilizer business and halt the appalling waste of farm land by soil erosion. Even more stimulating to chemically minded imaginations is the production of elemental phosphorus by a private company that not only has succeeded in extracting it from the phosphate rock in electric furnaces but actually has contrived to capture it and to transport it safely in tank cars. New values have been created for low-grade phosphate deposits, and the economic life of reserves has been lengthened while the waste of former years goes back into the mills for recovery of its values by modernized methods. Tennessee's phosphate-mining industry dates from the discovery of blue rock in Lewis County in December 1893. Its known reserves are of the order of 101,500,000 tons, which compares with the production of 825,099 tons in 1937; that is, 21 per cent of the phosphate-rock output of the whole nation. The total output reported from 1894 until July 1. 1938, was approximately 19,000,000 tons. Offcial estimates of reserves are overly conservative and evidence submitted before a joint Congressional committee in November 1938 places "probable" reserves at more than 6,000,000,000 tons.

Jan 1, 1942

By Robert Linton

In 1898, Joseph L. Giroux and J. A. Snedaker organized the Pilot Knob Copper Co. and began developing the Pilot Knob mine at Kim-berly, Nev., for high-grade copper ores, carrying good gold and silver values, that the formation had disclosed. At that time the nearest railroad was over 150 mi. distant. The high-grade ores did not prove to be extensive at this point—at least to the depth (240 ft.) to which the work was carried—but development of neighboring claims indicated extensive deposits of sulfide porphyry ore of concentrating grade lying at shallow depth, and deeper deposits of direct-smelting oxidized ores occurring in limestone near the limestone-porphyry contact. The properties passed through the hands of several organizations, with only meager development, until acquired in 1913 by Consolidated Coppermines Co., whose holdings now comprise 3480 acres of mining claims in the Robinson district. Since that time development has been prosecuted more actively; and for the purpose of determining the most efficient and economical method of treating the ores, considerable testing work has been done during the past three years, both in the laboratory and in actual mill operations. At first, the shipments were of ore high enough in grade to be profitable even with the long wagon hauls to the railroad and prevailing freight and smelting charges. In 1905, a blast furnace and a concentrator were erected at Kimberly. On account of insufficient ore supply, the blast furnace was never blown in and later was dismantled. The concentrator as originally equipped was operated for only about two months, both because the ore supply at that time was not sufficient and because gravity concentration did not yield satisfactory extraction. The location and arrangement of the mill further entailed high operating costs. In 1908, the Nevada Consolidated Copper Co. began to mill successfully, on a large scale, its sulfide porphyry ores, which are similar in character to the adjacent Coppermines ores. After that time, all of the ores and concentrates produced by Coppermines Company were smelted and a considerable part of the Coppermines low-grade ores concentrated at the Steptoe works of the Nevada Consolidated Co.

Jan 1, 1921

By Donald N. H. Chi, Henry E. Perlee

Accident cost data bases for all mining sectors covering the period 1975 through 1980 have been created on the PDP-11/70 and VAX-11/780 computers at the Bureau of Mines Pittsburgh Research Center. The total expected costs of the following tangible elements are computed: (a) loss in personal income; (b) compensation of wages from State, Federal, and union funds for disabling injuries; (c) benefits for injuries resulting in death or permanent disability; (d) medical treatment and hospital care; (e) investigation of a fatal accident; and (f) production loss due to accidents involving fatalities or amputation injuries. Each cost record computed contains a reference to the original record in the Health and Safety Analysis Center (HSAC) accident files to permit concurrent analysis. The intended applications include examinations of annual trends in mine accident statistics on the basis of economic loss. This provides an additional perspective on mine accidents. Various examples of the use of these data bases are presented.

Jan 1, 1983

By J. Wesley Pullman

It is stated by Dr. Tuttle in a paper read before the New Jersey Historical Society, that the celebrated Dickerson mine at Succasunna, Morris Co., N. J., yielded ore, about as early as 1710, for use at the Whippany forge, on Rockaway River, between Rockaway and Boonton. The ore was, for some time, free to all who chose to dig it. The tract embracing the Dickerson mine mas taken up, on account of its minerals, in 1713 or 1714, by John Reading, who sold it in 1716 to Joseph Kirkbride. John Jackson built a forge in 1722 on the Rockaway River a mile west of Dover. The ore here worked was probably chiefly from the Dickerson mine, but a portion may have been Hibernia. Hon. E. D. Halsey, of Rockaway, K. J., writes me: It is quite possible that the ore was at first taken from the outcrop of Hibernia by anybody who wanted it. The property first had an owner, May

Jan 1, 1886

The University of Chicago Press, 5750 Ellis Ave , Chicago, Ill This Press publishes books on a wide variety of subjects of scientific interest, as well as several journals Only a few books on geology have been published, however, and the Journal of Geology. The Journal of Geology: A semiquarterly standard American publication in the geological field. $6 per year, single copies, $1, back numbers, within the past five years, $6 per volume Books on geology: The origin of the earth, by T. C Chamberlin, $2 50, Two solar families, by T. C Chamberlin, $2 50, Quantitative classification of igneous rocks, by W. Cross, J. Iddings, L V Pirsson and H. S Washington, $1.75; Studies in minor folds, by C. E. Decker, $1 50; An introduction to the study of ore deposits, by F. H Hatch, $3, A descriptive petrography of the igneous rocks, by A. Johannsen, $4 50, Essentials for the microscopical determination of rock¬forming minerals and rocks, by A. Johannsen (revised ed ) $3, Outlines of geologic history, with special reference to North America, edited by R D. Salisbury, 8150

Jan 1, 1933

By Hertha R. Freche

The diffusion of magnesium and silicon from the core into the high-purity coating of Alclad sheet is important commercially, and led to the use of a duplex product for the study of diffusion by means of spectro-graphic analysis of successive layers. It has been found that in the duplex materials the diffusion of magncsium from an undersaturated aluminum-magnesium alloy into its high-purity aluminum coating is approximately of the same order of magnitude as the diffusion of silicon from a saturated aluminum-silicon alloy into its high-purity aluminum coating. However, if magnesium and silicon are both present in the alloy in the ratio of Mg2Si, they appear to diffuse in the approximate ratio of this compound. Finally, if the alloy contains silicon in excess of the amount required by magnesium silicide, the amount of diffused silicon in the high-purity aluminum coating will also be in excess. The interdiffusion of metals in the solid state is of both scientific and technical interest, because it is helpful in the study of the alloy structure and in the solution of problems pertaining to the strength and surface condition of the fabricated product. W. C. Roberts-Austenlt carried out the first quantitative determination of the coefficient of diffusion of gold into lead. This classical experiment was followed by a series of investigations on the diffusion of metals and nonmetals in iron and steel and in nonferrous metals and alloys. Recently the diffusion of copper in coated strong aluminum alloys was investigated by A. Burkhardt2. This appeared to be the first published investigation on the diffusion of metals into aluminum. The present investigation was the direct result of a technical question in connection with alloy fabrication and is a summary of some preliminary work pertaining to the diffusion of magnesium and silicon into aluminum. The results necessarily will be subject to some variations, since this is the first attempt to evaluate the coefficients of diffusion of magnesium and silicon into aluminum according to this method of procedure. This

Jan 1, 1936