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By W. C. Ellis, M. E. Fine

INTRODUCTION IN the iron-cobalt system there are several property-composition relationships of theoretical importance. The alloys are ferromagnetic exhibiting a maximum saturation at approximately 33 at. pct cobalt;1 the interatomic distances at room temperature show large positive deviations from Vegard's rule;2 and in the region of 50 at. pct cobalt there is an order-disorder transition which has a large effect on other properties. The thermal expansion data for the system have been reported3,4.5,6 only from room temperature to 100°C, and for this temperature range a minimum in coefficient in the region of 50 at. pct cobalt is well established. 3,4,6 This paper reports the expansion properties in the iron-cobalt system for temperatures from 30 to 850°C, shows the correlation with other physical properties, and proposes an explanation for the interrelation. EXPERIMENTAL PROCEDURE The thermal expansivities of iron-cobalt alloys in 10 pct steps in composition were determined from 30 to about 850°C. The major series of alloys was prepared by induction melting from Armco iron and cobalt rondelles with the addition of about 0.6 wt pct manganese. This series was supplemented by three compositions: high purity electrolytic iron, high purity electrolytic cobalt, and an alloy of 52.1 wt pct cobalt prepared from the high purity electrolytic iron and electrolytic cobalt by vacuum-melting. The compositions of both series of alloys are given in Table I. The specimens for measurement were prepared from the cast ingots by first hot-swaging followed in most instances by cold-swaging to 0.180-in. diam rod. Some alloys in the region of 50 pct cobalt, because of inherent cold-brittleness, were swaged hot to the final diameter. The expansion specimens were approximately 2 ½ in. long with flat-ground and polished ends. Before measurement they were annealed in hydrogen for r hr at 900°C. In the apparatus used for the thermal expansivity measurements (Fig I), a modification of that designed by Hidnert and Sweeney,7 the differential expansion between the specimen and vitreous silica was measured by a dial micrometer reading to 0.001 mm. An invar fixture supported the dial gauge, and this fixture, in turn, was connected to a split invar collar which supported the outer silica tube. A pyrex "T" was placed over the outer silica tube and was sealed to the invar collar by Apiezon Q sealing compound. A movable hollow vitreous silica rod which rested on top of the test specimen slid through a flexible rubber diaphragm cemented to the top of the pyrex "T." The specimen rested on a vitreous silica pedestal fused into the outer tube and was kept centered by a silica sleeve. A non-oxidizing atmosphere was obtained by admitting hydrogen into the

Jan 1, 1948

By W. E. Ruder

It has been pointed out by Stead1 that grains of considerable coarseness may be developed in steels containing from 3 to 5 per cent. of silicon, and in a previous paper2 the present author has shown that under the proper conditions of annealing single grains having an area of 50 sq. cm. may be obtained. It is the purpose of this paper to discuss in more detail the conditions promoting the growth of these grains, in the hope that the experiments described may throw some additional light upon the general problem of grain growth in metals. These silicon steels lend themselves particularly' well to such experiments because of the ease with which the change in structure may be followed and because of the fact that the effect of the carbon present is largely obliterated by the high percentage of silicon. This forms a solid solution with the iron, which acts very much like a pure metal. This alloy has been found to have but one critical point, viz., Ar2. This point, determined by Roberts-Austen for Hadfield's 4 per cent. Si alloy, was found to lie at 703" C. Vigouroux3, in working with pure Fe-Si alloys, finds this point to rise from 750" C. for 0.5 per cent. Si to 860" for 6 per cent. alloy, with a flat region in the curve from 3 to 4 per cent. Si. No accurate determination of this point has been undertaken by the present author, but it has been roughly determined to lie between 730" and 750:. In the preliminary experiments, some of which have been described in a previous paper, it was found that practically all of these steels, when annealed to 1,200" C. or higher, showed grains of comparatively large size. These grains varied in size, however, from about 4 sq. mm. to 50 sq. cm. (see Figs. 1 and 2) under the same conditions of anneal, and it was for the purpose of determining the cause of this difference in size that the experiments herein described were undertaken. From a preliminary study of the chemical composition of the specimens

Jan 1, 1914

By W. M. Weigel

Economic deposits of barite occur in Missouri in two main districts. The most important, the Southeastern or Washington County district, is in the southeastern part of the state, mostly in Washington County with a small extension into the northeastern part of St. Francois County. Roughly, it embraces nine townships or about 200,000 acres and accounts for from 80 to 90 per cent. of the state production. The Central district, is in the central part of the state, includes most of Morgan and Miller counties, and parts of Camden; Benton, Moniteau and Cole counties. The area embraced is much larger than the Washington County district, but the deposits are more infrequent. The location and general area covered by the two districts is shown in Fig. 1. There is another small area in southern Franklin County but so far its production has been of small consequence. It might be considered as a northward extension of the Washington County field. An, area in the central part of Hickory County has also produced a small amount. The outline of the districts as shown indicates the general area in which deposits have been found. The total area covered by deposits of barite is only a small part of this, as, in the Central district especially, there may be miles of supposedly barren areas between ore-bearing areas of an acre or two in extent. The date when barite was first mined in Missouri cannot be definit stated, but the industry probably dates back to about 1860. The occurrence of barite was known long before this, as it was encountered in the lead mines in Washington County and mentioned by several earlier writers." Missouri is reported as producing 8000 tons in 1882 so it is evident the industry was well under way in the seventies in order to have arrived at this substantial production. By 1893 Missouri production had increased so that it ranked with that of Virginia, then the largest producer it increased until in 1914 Missouri accounted for about 65 per cent. of the total United States output.

Jan 1, 1929

By C. Richard Tinsley

Like most minerals, coal is inherently a demand-limited commodity. The very sedimentary nature of its occurrence implies greater availability potential than demand. But this situation is overridden by economics among fuels, between coals, and within coal blends. Such considerations make coal forecasting a very hazardous profession indeed. THERMAL COAL If one thought that the lead times involved with a mining project were very long, one has obviously not been exposed to the planning process in the electric generation business - a process seriously confounded by shifts in load growth, environmental pressures, capital intensity, security of fuel sourcing, inter-fuel economics, and so on. But as a general rule, the near-term forecasts for thermal coal can reliably be based on a bottom-up, plant-by-plant analysis. Cement plant conversions can also be reasonably estimated next in order of reliability, although they have a much wider spectrum of coal qualities and fuel sources to choose from with a notably higher tolerance for sulfur and ash. Finally, industrial demand can be assembled from the estimates for conversions by pulp/paper plants, chemical plants, etc. The industrial sector is harder to estimate, since it may involve small boilers or dual-fired units. Assessing demand in the Pacific Rim is relatively a straightforward process in the near term because the major importing countries are all located on the Asian continent with either negligible or very minor (yet stable) indigenous coal production, (itself often operated on a subsidized basis). Furthermore, all imports are seaborne. These major importers are Japan, Korea, Taiwan, and Hong Kong with Thailand, Singapore, and Malaysia up-and-coming consumers. The suppliers to this market all have substantial reserves to back up decades of exports to these countries. Australia, the US, Canada, South Africa, China, and the USSR dominate the supply side. The second oil-shock of 1979/1980 has convinced the importers that reliance on oil can be expensive and eminently interruptible. Thus, they are determined to diversify away from oil' to nuclear and coal for generating electricity and for coal for other purposes where possible. This trend is seen to continue even in the face of the oil glut worldwide and oil-price reductions in early 1982. But the importers are also convinced that reliance on one coal source and, in particular, one infrastructure route for the coal chain from mine to consumer can be equally expensive and interruptible. Strikes in the US and Australia; excessive demurrage at certain ports; relegation of coal to a lower priority on multiple-use railroads in the USSR and China; and concern over escalation on high-infrastructure or high-freight coal chains are among the risks worrying the importers. As a consequence, Pacific Rim thermal coal purchases are being allocated among supplier nations, between ports, and within each country. An example of Japan's shift away from Australia and toward the US and Canada is shown in the estimates in Table 1. But the confidence of the import estimates deteriorates sharply beyond the plant conversion timetables and construction schedules in the near term. If part of the second generation of coal-fired power plants can handle lower-energy coals, the field of suppliers could widen to accept sizeable tonnages from Alaska, Wyoming, Alberta, or New Zealand resources. These supply sources generally have some infrastructure or freight advantage to compensate for their lower quality and to compete on a delivered energy-unit basis. These also offer diversification in sourcing. And the possibility of coal liquefaction in Japan further widens the sourcing network. A great number of Pacific Rim coal forecasts have been generated, especially for Japanese thermal-coal imports which are expected to grow strongly in the 1980's. Since the Japanese themselves have not yet settled their energy policy, the exact numbers are hard to call. Nevertheless, at 50 million tonnes of imports in 1990, Japan would consume 50-60% of the total Asian thermal coal imports as shown on Tables 2 and 6. The next most important consumers are the "island" nations of Korea, Taiwan, and Hong Kong (see Tables 3-5). All three are embarking on power plant developments usually with captive unloading facilities, capable of accepting more than 100,000-dwt vessels. Korea, with no-indigenous bituminous coal, is not especially enamoured with US coals, which are deemed too heavily loaded by freight and infrastructure costs -- up to 70% of the delivered price. Thermal coal contracts are presently split to Australia (70%) and to Canada (30%). Korea Electric Power Co. is already considering second-generation boilers capable of burning lower-quality coals than the present standard. Korea does burn domestic anthracite.

Jan 1, 1982

By Hisaaki Takata, Nobukazu Nanko, Ichiro Izawa

INTRODUCTION Tochibora Mine belonged to Kamioka Mines, which are located in Gifu Prefecture, the cen- tral part of the Japanese main island, has occupied the important position as a main source of silver, lead and zinc for the smelter of Mitsui Mining & Smelting Co., Ltd.. The annual production of crude ore goes up to about 1 mil- lion metric tons with average grades of 20g/t Ag. 0.31% Pb, and 4.18% Zn. For the last 20 years, the mine has struggled with the depressed zinc metal price in yen and the escalation of operating cost, especially la- bour cost. For instance, prices of commodities and labour cost have been raised 8 times and 14 times respectively during this period. On the other hand, total operating cost has been con- trolled to increase 4.5 times. The productivity of the mine has been increased 4 times by means of improvement of individual works and adoption of perfectly mechanized trackless mining system to reform the mining method. Some improvements are presented as follows; Today, drifting is carried out by the tunnelling crew which consists of 9 operators equipped with 5 Mobile-Jumbos, 2 LHD's and a jeep for carrying and loading ex- plosives. In order to get the high utilization of these machines, each crew manages systematically and continuously adjacent 15 faces under the direction of a forceman. As a result, we have accomplished high advance of drift. Furthermore, raising is carried out by a set of the raise-borer in case of long raises like ore- passes and also by the stage blasting method in case of short ones like slot-cuts. Generally speaking, the sublevel stoping method is considered to be one of the best methods for various respects; high productivity, low cost, easy grade control and safety. This method has been introduced in Tochibora Mine since 1950 and is now widely applied to lodes that could not be mined before by this method be- cause of the shapes of them and the conditions of their country rock. Now we are convinced to be nearly accomplished for our method. This report presents some examples of the standard sublevel stoping method and its variations which are practised at the mine. STANDARD SUBLEVEL STOPING Up to 1979 since 1950 when this method was first applied at the mine, 17 million tons of crude ore have been produced by this method, which amounted to 55% of the total production of the mine. During three decades, the productivity has been increased considerably by means of trackless mining system using large-sized machines and etc.. In other words, mining efficiency has been increased from 20 tons/man.shift to 100 tons/man.shift at large stopes. Formerly undercutting was carried out above the drawing drift by jack-legs and small scrapers and then several cone-shaped glory holes were made by stopers between the undercut level and the draw level. But now, a undercut drift is driven in parallel with the haulage drift at the same level of it. And at the undercut drift, drilling upward is carried out by a large-sized machines like the fan drill in order to extract the broken ore directly at the drawing drift connected with the haulage drift. Simplification of the work as well as utilization of large- sized machines enables to get higher productivity. But if the deposit is comparatively wider and the broken ore around the middle part of the deposit cannot be extracted, several drawing drifts are driven to the deposit and in order to minimize dead ore at the final stage of operation, the

Jan 1, 1981

By H. B. Pulsifer

THIS paper outlines a method for rapid production of flat, granular surfaces on many of the medium hard- alloys. Grinding wheels and fabrics on wheels are not used; the purpose is accomplished by rubbing on emery papers and on fine abrasives heaped on hand boards. The advantages of chromic acid in the solutions for etching nickel and copper alloys are explained. Steels may require an additional step in the mechanical smoothing owing to the lack of an ideal etchant. Chloric acid is a. rapid etchant for steel. Thirty-six photomicrographs illustrate the steps in smoothing and the clearness with which structures can be rapidly exposed. New structures in low-carbon steels are announced. After a little practice, many materials can be finished within two or three minutes. CUPRONICKELS In the previous papers on the metallography of copper and magnesium which the author has presented to this Institute there has been an attempt to demonstrate the advantages of minimizing the mechanical work of surfacing and depending more on chemical means; the present paper continues the research into a still wider field. It is obvious that the success of a chemical method for smoothing and etching depends almost entirely on the use of a suitable reagent. The etchant must remove the surface granulations and outline the structures. A very special type of reaction between reagent and metal must be displayed. Promiscuous corrosion or the intensification of the attack about local couples -formed by impurities or components must not be in evidence. There are several reagents that etch highly polished nickel and copper alloys very well. Acidified ferric chloride solutions, concentrated and dilute nitric acid solutions, and nitric and acetic acids in acetone have -been used to good advantage for the purpose. However, if they are expected to etch rather deeply-that is, smooth as well as characterize-the results may be far from satisfactory.

Jan 1, 1928

With a total value to date of well over $200,000.000.00 for its ore production, the Tintic mining district, which is about 100 miles southwest of Salt Lake City, ranks as one of the three main ore producing areas of Utah. In it are located the biggest silver mines of the country, silver having been the chief metal produced in the camp since its start. Tintic's silver production to date has a total value of over $1 00,000,000, approximately half of the value of the camp's ore production, the balance being roughly divided among lead, gold and copper. Discovery of the camp was made in the late sixties. Lack of transportation and ore reduction facilities delayed extensive development of the area for nearly a decade. In recent years Tintic's ore production has ranged from 350,000 to 400,000 tons annually. In 1924 the camp produced one-eighth of the silver produced in the nation with an output of 7,729,700 ounces. Its lead production was 77,765,000 pounds and its gold production 26,285 ounces.

Jan 1, 1925

By B. J. Nelson, F. N. Rhinos

Studies upon the rates of oxidation of copper alloys containing small quantities of the alloying elementsl,2 have shown that steady growth of the scales at predictable rates is limited to a small concentration range and to oxidation at temperatures above 600°C., whereas unpredictable and erratic oxidation behavior is found with alloys containing larger additions and at low temperatures. The latter effect appears to be associated with modifications in the geometric disposition of the oxides. At low concentrations of the baser elements in copper, simple oxidation behavior is typified by the formation of two distinguishable zones of oxidation: (I) a subscale composed of discrete particles of the oxide of the alloying element embedded in substantially pure copper and (2) an external scale composed of cuprous oxide (with a very thin layer of cupric oxide on the outside) containing particles of the alloying element distributed throughout, but most profusely near the metal surface (see Fig. 2, top). Where this type of structure obtains, the process of oxidation is believed to be implemented, and its rate controlled, by the diffusion of the reacting elements through a' continuous matrix phase; i.e., through the alpha (metallic) phase of the subscale and alloy and through the cuprous oxide of the external scale. Copper is presumed to diffuse outward through the cuprous oxide layer to the external surface, where it reacts to form more cuprous oxide.3 Oxygen is delivered at the oxide-metal interface in the form of cuprous oxide, which may either react with the alloying element to deposit its oxide at the inner limit of the external scale or dissolve in the pure copper that remains at the surface of the metal after it has become depleted in its oxidiza-ble alloying elements. Both oxygen and the alloying element are pictured as diffusing through the metal, the oxygen inward and the alloying element outward. Where the two diffusion streams meet in sufficient concentration the oxide of the subscale is precipitated. If the oxygen pressure in the surrounding atmosphere is just below the pressure necessary to maintain cuprous oxide undecomposed, no cuprous oxide can be formed; some of the oxide of the alloying element may form upon the external surface, if its decomposition pressure is below that of cuprous oxide, and a normal sub-scale appears. This condition is achieved experimentally by enclosing the alloy with cuprous oxide together with an excess of copper. At temperatures from 600°C. downward, the precipitated oxides tend to accumulate at the grain boundaries of the alpha phase where a relatively small volume establishes a partial barrier to the diffusion of the reactants- Similarly, large volumes of Precipitated oxides tend to deposit in various patterns that may be expected to

Jan 1, 1944

By F. N. Rhinos, B. J. Nelson

Studies upon the rates of oxidation of copper alloys containing small quantities of the alloying elementsl,2 have shown that steady growth of the scales at predictable rates is limited to a small concentration range and to oxidation at temperatures above 600°C., whereas unpredictable and erratic oxidation behavior is found with alloys containing larger additions and at low temperatures. The latter effect appears to be associated with modifications in the geometric disposition of the oxides. At low concentrations of the baser elements in copper, simple oxidation behavior is typified by the formation of two distinguishable zones of oxidation: (I) a subscale composed of discrete particles of the oxide of the alloying element embedded in substantially pure copper and (2) an external scale composed of cuprous oxide (with a very thin layer of cupric oxide on the outside) containing particles of the alloying element distributed throughout, but most profusely near the metal surface (see Fig. 2, top). Where this type of structure obtains, the process of oxidation is believed to be implemented, and its rate controlled, by the diffusion of the reacting elements through a' continuous matrix phase; i.e., through the alpha (metallic) phase of the subscale and alloy and through the cuprous oxide of the external scale. Copper is presumed to diffuse outward through the cuprous oxide layer to the external surface, where it reacts to form more cuprous oxide.3 Oxygen is delivered at the oxide-metal interface in the form of cuprous oxide, which may either react with the alloying element to deposit its oxide at the inner limit of the external scale or dissolve in the pure copper that remains at the surface of the metal after it has become depleted in its oxidiza-ble alloying elements. Both oxygen and the alloying element are pictured as diffusing through the metal, the oxygen inward and the alloying element outward. Where the two diffusion streams meet in sufficient concentration the oxide of the subscale is precipitated. If the oxygen pressure in the surrounding atmosphere is just below the pressure necessary to maintain cuprous oxide undecomposed, no cuprous oxide can be formed; some of the oxide of the alloying element may form upon the external surface, if its decomposition pressure is below that of cuprous oxide, and a normal sub-scale appears. This condition is achieved experimentally by enclosing the alloy with cuprous oxide together with an excess of copper. At temperatures from 600°C. downward, the precipitated oxides tend to accumulate at the grain boundaries of the alpha phase where a relatively small volume establishes a partial barrier to the diffusion of the reactants- Similarly, large volumes of Precipitated oxides tend to deposit in various patterns that may be expected to

Jan 1, 1944

THE shipments of high-grade manganese ore, con-taining 35 per cent or more of manganese, from the mines in the United States in 1926 were slightly less than half as large as similar shipments in 1925, accord-ing to the United States Bureau of Mines, Department of Commerce. The shipments in 1926 by 45 producers amounted to 46,258 long tons as compared with 98,324 long tons by 42 producers in 1925. This decrease was due to the falling off of the shipments from Montana. The domestic production of metallurgical ore in 1926 was 26,530 tons; that of chemical ore, 19,728 tons. Montana was the largest shipper with 5713 tons of metallurgical ore and 17,904 tons of chemical ore. The largest individual shipper of metallurgical ore was the Crescent Mine in the Olympic Mountain Region, Clallam Co., Washington, from which the shipments for the year were 3162 tons. The mine was operated by lessees and reverted to the owners in July, and thereafter to the close of the year no .ore was shipped.

Jan 6, 1927

McCulloch, Leon.-Experiments with Sherardizing. Discussed by Fred. L. Wolf, Willis M. Peirce, Jesse L. Jones, O. W. Storey, David R. Kellogg, William H. Finkeldey, L. H. Marshall 2 Bassett, W. H. and Davis, C. H.-Spectrum Analysis in an Industrial Laboratory. Discussed by C. S. Brainin, C. H. Davis, W. H. Bassett, George K. Burgess, George K. Elliott, Leonard Waldo, Keivin Burns 8 Roast, Harold J. and Pascoe, Charles F.-Arsenical Bearing Metals. Discussed by Jesse L. Jones, William A. Cowan,. William K. Frank, William M. Corse, Harold J. Roast 10 Bain, Edgar C.-Crystal Structure of Solid Solutions. Discussed by Ancel St. John, Edgar C. Bain, Paul D. Merica 14 Peirce, Willis M.-Studies on the Constitution of Binary Zinc-base Alloys. Discussed by George K. Burgess, Willis M. Peirce, D. H. Ingall, J. L. Haughton, K. E. Bingham 16 Webster, William R.-Application in Rolling of Effects of Carbon, Phosphorus and Manganese on Mechanical Properties of Steel. Discussed by George B. Waterhouse, J. O. Liebig, Ralph H. Sweetser, H. T. Morris, F. D. Carney, W. P. Barba, John H. Nelson, George K. Burgess 19 Barba, W. P. and Howe, Henry M.-Acid Open-hearth Process for Manufacture of Gun Steels and Fine Steels. Discussed by W. J. Priestley, Haakon Styri, C. T. Patterson, Harry Brearley, W. P. Barba, George F. Comstock, Alexander L. Feild 24 Kellogg, David R.-Electrolytic Deposition of Iron for Building up Worn or Undersized Parts. Discussed by Jesse L. Jones, William Blum, C. R. Hayward, George K. Burgess, C. P. Perin, David R. Kellogg 33 McQuaid, H. W. and Elm, E. W.-Effect of Quality of Steel on Case-carburizing Results. Discussed by H. J. French and L. Jordan, W. J. Priestley, M. H. Schmid, A. E. White, H. J. Stagg and S. C. Spalding, Henry S. Rawdon, J. S. Vanick 35 Phillips, Arthur and Davenport, E. S.-Malleableizing of White Cast Iron. Discussed by Samuel Epstein, H. A. Schwartz, A. E. White, R. S. Archer.... 58 Priestley, William J.-Effect of Sulfur and Oxides in Ordnance Steel. Discussed by Harry T. Morris, William J. Priestley, C. P. Perin, James H. Gray, W. P. Barba, Haakon Styri, Frank D. Carney, W. L. Harbrecht, George F. Comstock, E. J. Janitzky 67

Jan 5, 1922

No true conception of the Park City mining district can be obtained without first giving consideration to the part it has played as a consistent producer of mineral wealth. Its position in this regard is almost unique in the annals of a lead silver camp. For over fifty years it has been a continuous producer of lead and silver and for twenty years it has added to the world's supply of zinc. Few camps with its productive history are so little known to the mining fraternity for the reason that its life has never been punctuated by periods of "boom" which serve to attract general attention, neither has it been the field of operation for promoters who temporarily focus attention upon their scene of action. Its history is one of legitimate mining operations. These were at first carried on by a number of smaller companies. A period of consolidation followed and at the present time the entire production of the camp is in the hands of two major companies and one smaller company. These are flanked by a number of prospecting properties, which are at this time actively engaged in the search for ores and later events will doubtless put some of them in the producing class.

Jan 1, 1925

By A. W. Schlechten, R. F. Doelling

Parkes' process crusts were vacuum distilled using a shortened Pidgeon retort. Zinc was effectively removed below 800°C and recovered as a zinc sheet easily stripped from the furnace liner. Lead dasdistillationazinc required about 950°C and the resulting lead condensate tillationtended to stick to the thin metal liner. Final purity of residue is limited only by nonvolatile metals such as copper. THE successful commercial application of vacuum distillation for the removal of zinc from lead, as described by W. T. Isbell,' raises the question of whether or not a similar method can be used for other metal separations. Preliminary laboratory experiments on the vacuum distillation of Parkes' crusts by one of the authors' gave such encouraging results that it was decided to make a series of tests with large scale equipment to determine the results that might be expected in plant operation. The pilot-plant test results are reported in this paper. The great advantages of vacuum dezincing are its simplicity, the relatively low temperature of operation, and the fact that the zinc is recovered in the metallic state ready for use again in the desilveriza-tion of lead bullion. These same advantages and others would be realized from the vacuum distillation of the zinc-lead-silver crusts made in the Parkes' process. Zinc readily will distill over at temperatures of 800" to 900 °C as contrasted with the 1200°C or more used in the Faber du Faur furnace. Furthermore, the recovery of zinc as metal is practically perfect since there is no loss by oxidation or fuming. If the vacuum distillation is continued and hastened by raising the temperature to around 1000°C, the lead will distill over at an appreciable rate and deposit in a warmer zone of the condenser. This lead will contain a small amount of silver, but it can be returned to the desilverizing kettles and there is not the chance of loss that there is in disposing of the rich litharge from cupelling. The lower temperatures required will mean a saving of fuel and less wear on the equipment. One further point is that the vacuum distillation could be run on a small scale practically continuously, so as to avoid the accumulation of large amounts of silver-rich material and the need of assembling a large crew for special cupellation runs. It should be pointed out here that the final purity of the silver residue will depend on the copper or iron content of the crusts, as these and other metals with low vapor pressure will not be eliminated. The idea of treating the silver-zinc-lead crusts from the desilverization of lead bullion by vacuum distillation is not a new one, but such experiments have not been reported, as far as can be determined. In 1912, T. Turner presented a paper3 before the Institute of Metals on the vacuum distillation of brass, and a member of the audience suggested that he try his method on Parkes' crusts. In 1935, W. J. Kroll, noted for his discovery of debismuthiz-ing lead by calcium additions, published a paper' on vacuum distillation of metals and pointed out possible applications in the lead industry. Principles of Separation The successful separation of two or more metals by distillation depends upon an appreciable difference in the effective vapor pressures of the metals in the alloy to be treated. The vapor pressures of the common metals have been determined with considerable accuracy; for example, at 800°C the vapor pressures of the individual metals which are of particular interest in this paper are about .as follows: Zn, 241 mm; Pb, 0.0553 mm; and Ag, 0.0000276 mm. Unfortunately,, data are not available on the activities of the components in the ternary system Zn-Pb-Ag, but it can be estimated that on the basis of the vapor pressures of the individual metals a good separation can be made by distillation. The preliminary laboratory tests previously mentioned' supported this assumption. The pilot-plant size vacuum distilling furnace is shown in Fig. 1. It consisted of an abbreviated Pidgeon retort, 5 ft long with an ID of 8 in., heated by six silicon carbide elements placed symmetrically

Jan 1, 1952

By C. A. Hartnagel, D. H. Newland

The market for Pennsylvania grade showed a drop both in demand and in prices, which reacted unfavorably on New York oil production, all of which comes under that classification. The output for the year in New York amounted to 5,045,200 bbl. The production in 1937 was 5,478,000 bbl., the largest in any year for the last half century or more. Falling prices, of course, affected new drilling. They declined from $2.20 a barrel, which was posted at the start, to $1.68 in the final months of 1938. With this came a reduction in the amount of oil that would be accepted by the refineries operating in the New York fields. Conditions therefore have combined to halt, at least temporarily, the growth of oil production in the state, which had experienced a steady gain for the last 15 years or more, the yield having risen from less than 1,000,000 bbl. to the round total of 5,500,000 bbl. in 1937. Practically all the oil now obtained is from old, developed fields. No new pools have been found in the last quarter of a century or more. The larger share comes from flooding the sands. Although there are many wells that still produce without the aid of the water drive, they are small and for the most part are pumped periodically. Of late, new drilling has been restricted mainly to that required to maintain existing flood projects. Because of the high initial cost and time required to put a flood into production, the depressed condition of the market has been a serious handicap financially, and, further, has been an unfavorable influence on the planning of operations to secure the maximum recovery of the oil. Unless market conditions improve, it is certain that the effects will be shown in the yield for some time to come. Gas Developments Drilling was active in the southwestern area, chiefly in Allegany and Steuben Counties, and some notable wells were brought in that tapped the Oriskany sandstone. Altogether 42 holes were put down in the two counties, of which 31 were listed as producers and 11 were dry. The new wells gave an initial daily flow of 239,000 M cu. ft. At the close of

Jan 1, 1939

By C. A. Hartnagel, D. H. Newland

The market for Pennsylvania grade showed a drop both in demand and in prices, which reacted unfavorably on New York oil production, all of which comes under that classification. The output for the year in New York amounted to 5,045,200 bbl. The production in 1937 was 5,478,000 bbl., the largest in any year for the last half century or more. Falling prices, of course, affected new drilling. They declined from $2.20 a barrel, which was posted at the start, to $1.68 in the final months of 1938. With this came a reduction in the amount of oil that would be accepted by the refineries operating in the New York fields. Conditions therefore have combined to halt, at least temporarily, the growth of oil production in the state, which had experienced a steady gain for the last 15 years or more, the yield having risen from less than 1,000,000 bbl. to the round total of 5,500,000 bbl. in 1937. Practically all the oil now obtained is from old, developed fields. No new pools have been found in the last quarter of a century or more. The larger share comes from flooding the sands. Although there are many wells that still produce without the aid of the water drive, they are small and for the most part are pumped periodically. Of late, new drilling has been restricted mainly to that required to maintain existing flood projects. Because of the high initial cost and time required to put a flood into production, the depressed condition of the market has been a serious handicap financially, and, further, has been an unfavorable influence on the planning of operations to secure the maximum recovery of the oil. Unless market conditions improve, it is certain that the effects will be shown in the yield for some time to come. Gas Developments Drilling was active in the southwestern area, chiefly in Allegany and Steuben Counties, and some notable wells were brought in that tapped the Oriskany sandstone. Altogether 42 holes were put down in the two counties, of which 31 were listed as producers and 11 were dry. The new wells gave an initial daily flow of 239,000 M cu. ft. At the close of

Jan 1, 1939

By R. F. Mehl, C. S. Barrett, A. G. Guy

INTRODUCTION IN the last few years this laboratory has published a series of papers on the mecha¬nism of age -hardening.14,1,6,11,20 Briefly stated it has been proposed that hardening is caused by lattice strains created by lat¬tice coherency between matrix and precipi¬tate; much experimental evidence has been advanced to support this thesis. The system Cu-Be is an attractive one to study. It is a system that exhibits marked age hardening; it ought to be subject to the type of analysis of structural changes hitherto explored and it is a prominent example of "discontinuous precipitation," which, though studied in other cases, 6,20 remains largely unexplained. This research shows that hardening in the Cu-Be system proceeds essentially as proposed earlier.14 Guinier-Preston zones have been observed to form on the {100) planes of the matrix; no transition lattice was found. Aging and overaging were ob¬served to occur most rapidly in the region of discontinuous precipitation-the grain boundary-apparently accelerated there by strain. The Constitution of the Cu-Be System The Cu-Be constitution diagram is shown in Fig I. The structure of the ? phase was determined by Misch16 as of the caesium-chloride type with a lattice con¬stant of 2.598 Å High temperature X ray photograms of 7.2 pct Be alloy made by Kossolapow and Trapesnikow13 led them to conclude that the structure of the P phase is simply disordered body-centered cubic with [ao] = 2.79 Å at 750°C. Previous Work The initial work on the age-hardening of Cu-Be alloys was that of Masing, Dahl, Holm, and Haase. Their numerous papers were published in book form in 1929, sub¬sequently translated.18 The effects of Be content, solution temperature, aging tem¬perature, and cold work on the hardness developed in these alloys were studied. An initial decrease in the electrical conduc¬tivity was observed when strips of a 2.5 pct Be alloy were aged at temperatures from 200 to 350°C; a wire of the same analysis failed to exhibit this decrease when aged at 350°C. A volume decrease of 0.6 pct oc¬curred during aging at 250°C. Microscopic investigations of quenched and aged alloys disclosed a regularly ori¬ented "striping" (ripples) on the a and on the ß grains. In the case of the ß the cause of the striping was held to be decomposition

Jan 1, 1948

By E. E. Donath

Application of the coal hydrogenation process for the production of chemicals is described. It has been estimated that a plant to produce 31,090 bbl per day of chemicals and fuels would cost $326,- 000,000. These products would consist of 6.1 pct by weight tar acids, 50.8 pct aroma tics, 16.4 pct LPG, and 26.7 pct gasoline. THE coal hydrogenation process is well known as a means for the production of liquid fuels from coal. In this paper the possibilities of the coal hydrogenation process as a source of chemical raw materials will be discussed. The recovery of chemicals from coal hydrogenation products necessitates some modifications and additions to the equipment used for the production of fuels. The conversion of coal into liquid fuels such as gasoline requires two steps: liquid phase and vapor phase hydrogenation. Intermediate oils obtained in the liquid phase contain oxygen and nitrogen compounds, especially aromatic compounds, since the coal itself is composed mainly of cyclic compounds. These oils are the raw material for the separation of high molecular weight aro-matics and of tar acids and tar bases. The relative amounts of heavy and middle oil—boiling above and below, e.g., 325°C—can be influenced by temperature, residence time, and catalyst of the liquid phase hydrogenation. In the vapor phase hydrogenation a solid bed catalyst is used to convert the intermediate oil from the liquid phase operation into hydrocarbons of the gasoline boiling range. Essentially pure hydrocarbons with a low content of sulphur, oxygen, nitrogen, and olefines are obtained. Depending mainly on the catalyst, the high ring compound content of the liquid phase oil can be preserved and gasoline rich in aromatics obtained. However, with other catalysts, these ring compounds can be split into gasoline with a high content of isoparaffins.' Again, depending on the operating conditions, products of varying boiling range can be obtained. Selection of catalyst and operating conditions therefore makes possible the production of different grades of motor fuel and also variations in the relative amounts of benzene, toluene, and xylene when aromatic hydrocarbons are produced. The schematic flowsheet, Fig. 1, shows the main steps of the coal hydrogenation process, and the product streams used for the recovery of chemicals are indicated. In the liquid phase hydrogenation stage, coal is made into a paste with recycle heavy oil. The coal paste is mixed with hydrogen and preheated at a pressure of 300 to 700 atm in heat exchangers and preheaters. The coal paste and hydrogen enter high pressure converters where the liquefaction of coal, conversion of the high boiling liquid materials initially obtained, into intermediate oils is performed at a temperature around 470°C. The reaction products enter the hot catchpot in which lower boiling oil vapors are separated from the residue containing unreacted coal, ash, and heavy oil. This residue, usually called H.O.L.D. (heavy oil let down), is processed to separate solid and liquid material. By thermal treatment of the residue, e.g., a coking process, distillate oil is obtained, which can be recycled as part of the pasting oil. This stream is rich in polynuclear aromatics and can be, and has been, used for the recovery of such material. The oil vapors and hydrogen from the hot catchpot are cooled in heat exchangers and a cooler, and the condensed oil is separated from the recycle hy-

Jan 1, 1953

By E. E. Donath

Application of the coal hydrogenation process for the production of chemicals is described. It has been estimated that a plant to produce 31,090 bbl per day of chemicals and fuels would cost $326,- 000,000. These products would consist of 6.1 pct by weight tar acids, 50.8 pct aroma tics, 16.4 pct LPG, and 26.7 pct gasoline. THE coal hydrogenation process is well known as a means for the production of liquid fuels from coal. In this paper the possibilities of the coal hydrogenation process as a source of chemical raw materials will be discussed. The recovery of chemicals from coal hydrogenation products necessitates some modifications and additions to the equipment used for the production of fuels. The conversion of coal into liquid fuels such as gasoline requires two steps: liquid phase and vapor phase hydrogenation. Intermediate oils obtained in the liquid phase contain oxygen and nitrogen compounds, especially aromatic compounds, since the coal itself is composed mainly of cyclic compounds. These oils are the raw material for the separation of high molecular weight aro-matics and of tar acids and tar bases. The relative amounts of heavy and middle oil—boiling above and below, e.g., 325°C—can be influenced by temperature, residence time, and catalyst of the liquid phase hydrogenation. In the vapor phase hydrogenation a solid bed catalyst is used to convert the intermediate oil from the liquid phase operation into hydrocarbons of the gasoline boiling range. Essentially pure hydrocarbons with a low content of sulphur, oxygen, nitrogen, and olefines are obtained. Depending mainly on the catalyst, the high ring compound content of the liquid phase oil can be preserved and gasoline rich in aromatics obtained. However, with other catalysts, these ring compounds can be split into gasoline with a high content of isoparaffins.' Again, depending on the operating conditions, products of varying boiling range can be obtained. Selection of catalyst and operating conditions therefore makes possible the production of different grades of motor fuel and also variations in the relative amounts of benzene, toluene, and xylene when aromatic hydrocarbons are produced. The schematic flowsheet, Fig. 1, shows the main steps of the coal hydrogenation process, and the product streams used for the recovery of chemicals are indicated. In the liquid phase hydrogenation stage, coal is made into a paste with recycle heavy oil. The coal paste is mixed with hydrogen and preheated at a pressure of 300 to 700 atm in heat exchangers and preheaters. The coal paste and hydrogen enter high pressure converters where the liquefaction of coal, conversion of the high boiling liquid materials initially obtained, into intermediate oils is performed at a temperature around 470°C. The reaction products enter the hot catchpot in which lower boiling oil vapors are separated from the residue containing unreacted coal, ash, and heavy oil. This residue, usually called H.O.L.D. (heavy oil let down), is processed to separate solid and liquid material. By thermal treatment of the residue, e.g., a coking process, distillate oil is obtained, which can be recycled as part of the pasting oil. This stream is rich in polynuclear aromatics and can be, and has been, used for the recovery of such material. The oil vapors and hydrogen from the hot catchpot are cooled in heat exchangers and a cooler, and the condensed oil is separated from the recycle hy-

Jan 1, 1953

By N. J. Sather

A detailed description is given of Bunker Hill's concentration process employed at the company's lead-zinc property in the Coeur d'Alene district, Idaho. The plant is equipped to process 3000 tpd of Pb-Zn ore. The history of the Bunker Hill mine dates back to Aug. 26, 1885, when Noah S. Kellogg found the outcrop of the Bunker Hill orebody on the hillside of Milo Gulch above the present town of Wardner, Idaho. A small concentrator was erected in 1886, and before the railroad came, the mine output was transported by mule teams to the head of navigation on Coeur d'Alene Lake. After ten years of production, an interest was purchased in the Tacoma smelter at Tacoma, Wash., to which the entire output was shipped. In 1916 and 1917, a modern lead smelting plant and refinery was erected, and a modern electrolytic zinc reduction plant was built in 1927-1928 by the Sullivan Mining Co. which was 50 pct owned (and now totally owned) by the Bunker Hill Co. The first concentrator, a 100-tpd capacity plant, was built in 1886 near the Reed Tunnel in Wardner. The treatment consisted simply of crushing ore to 2 in. and then to 1/2 in. through rolls in closed circuit with a bucket elevator and trommel screen. The undersize was further sized in trommels to 10 mm and 3 mm. Oversize from the latter was treated in Hartz jigs. The 3 mm undersize was classified in crude hydraulic classifiers furnishing sands to sand jigs, the overflow being dewatered in a long V tank which supplied sands to a Cornish buddle and to Frue vanners. The concentrates were of excellent grade, old records showing assays of 69 pct lead and 29 oz. of silver per ton. Tailings were high inasmuch as the provision for treating the slimes was inadequate. Since the erection of the first mill, the Bunker Hill plant has gone through many changes and improvements. In succeeding years, the flowsheet was enlarged somewhat with Huntington and Chilean mills, vanners, tables, ball mills, and flotation cells being progressively added. In the transition period to the present day practice, the object was not only to improve the equipment for better mechanical performance, but to add new equipment whenever practicable so as to improve both grade and recovery of the product at a lower cost. For the modern plant and flow-sheet, see Figs. 1 and 2. The history of flotation at Bunker Hill dates back to 1913 when experimental work was done in a cell in which the pulp was circulated through a centrifugal pump set outside the cell and discharged against a baffle in the cell. The cell was built with a spitzkasten, and the quiet zone in the latter tended to let the froth break down before it got to the overflow weir. To remedy this situation, a belt drag with light metal flights was arranged for drawing off the froth over a ramp in each cell. Recovery of lead was not as high as present day practice, but was so much better than the vanners that the cells were multiplied gradually to eight cell units. The mill men called them mud hens. Flow between the cells was controlled by a weir, which determined the pulp level in each cell. These early machines were called Bunker Hill machines. The tailings were later treated in pneumatic scavenger flotation machines in a series, producing concentrates for retreatment in the Bunker Hill cleaners as well as the final tailings. This system of flotation served with minor changes until the Fahrenwald machine was developed at Bunker Hill during the period from 1921 to 1925. In 1925, the Fahrenwald flotation machine gradually replaced the Bunker Hill machine, although the Hearing pneumatic scavengers for retreatment of the tailings from the Fahrenwald machines were retained in the lead section until 1928. The use of the Hearing pneumatic machine for cleaning the rougher concentrates was begun in 1924. In 1941, plans were made for remodeling and modernizing the concentrator, the goal being a modern concentrator employing the latest equipment and milling practices. The old wooden building was to be replaced by a steel and concrete structure. Shortly after preliminary work got under way, the United States entered World War II and the demands for metals increased. To meet this requirement, the management decided to increase the capacity of the concentrator to 1800 tpd. This was accomplished in record time by the construction and operation of a sink-float plant. Installation of this plant was the first step in the general plan for modernization. In 1947 the concentrator (including the sink-float plant) was enlarged to handle 3000 tpd. Milling operations today are at the rate of 2300 tpd, five days per week. During 1959, 453,000 tons of ore was beneficiated. The Bunker Hill concentrator is in three sections: the crushing plant, a 3000-ton capacity storage bin, and the concentrator building. The crushing plant is of wood design with a gunnite coating on the exterior.

Jan 1, 1961

By William H. Bassett, C. J. Snyder

Cable sheaths made from common lead have been used to protect underground cables since the beginning of the power-cable industry. Only within the last few years, however, has a systematic study of the production methods and of the causes of failures in cable sheathing been undertaken. Recent increases in the use and demands for service made upon power cables have emphasized the importance of high-quality cable sheaths. It was customary for the users to attribute the cause for their sheath failures to the lack of purity of the lead, but this contention has not been borne out by field and laboratory inspection of the sheaths. During the past two years the writers have had an opportunity to examine various types of sheaths that were removed from power cables after failure. These samples were obtained from various parts of the country and represent, it is supposed, the production of the principal cable manufacturers. Aside from corrosion, the common field difficulties can be classified either as mechanical or structural, as outlined below, and when structural irregularities are found they can all be attributed to manufacturing causes and in no way reflect the purity of the lead. Clasxification of Cable Sheath Failures Other Than Corrosive Mechanical Irregularities A. Splits caused by excessive internal pressure B. Injuries such as nail holes, cuts, tsars, etc. C. Intercrystalline cracks caused by vibration Structural Irregularities A. Splits C. Voids in wall of sheath a. Caused by imperfect welds a. Blisters 13. Holes through sheath b. Laminations a. Caused by yellow oxide inclu- D. Questionable structures sions in sheath wall a. Prominent flow lines b. Blowouts by entrapped gases b. Large areas filled with suboxide during manufacture c. Scattered yellow oxide or dirt C. Caused by inclusions of dirt and

Jan 1, 1933