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By Alden B. Greninger

The study of mosaic structure of crystals1 has been confined until recently to the field of theoretical physics. Crystallographers, in general, have neglected the subject, although X-ray crystallographers constantly make use of formulas for intensities of X-ray reflection as modified by mosaic structure. Work on mosaic structure has centered chiefly on the task of proving or disproving the secondary-structure theory of Zwicky,² and very little attention has been paid to the visible imperfections of crystals. A. Goetz has been responsible for most of the supposed experimental confirmations of Zwicky's theory.3 On the other hand, Smckal,4 and Orowan,5 among others, maintain that there is neither theoretical nor experimental evidence to substantiate Zwicky's secondary structure. Lately the subject has been intensively investigated by Buerger,6 one of whose recent papers led to a symposium on mosaic structures.'

Jan 1, 1935

By Alden B. Greninger

The study of mosaic structure of crystals1 has been confined until recently to the field of theoretical physics. Crystallographers, in general, have neglected the subject, although X-ray crystallographers constantly make use of formulas for intensities of X-ray reflection as modified by mosaic structure. Work on mosaic structure has centered chiefly on the task of proving or disproving the secondary-structure theory of Zwicky,² and very little attention has been paid to the visible imperfections of crystals. A. Goetz has been responsible for most of the supposed experimental confirmations of Zwicky's theory.3 On the other hand, Smckal,4 and Orowan,5 among others, maintain that there is neither theoretical nor experimental evidence to substantiate Zwicky's secondary structure. Lately the subject has been intensively investigated by Buerger,6 one of whose recent papers led to a symposium on mosaic structures.'

Jan 1, 1935

ALABAMA Anniston. Carrington, F. G. Ashland.-Herr, I. Auburn.-Brown, R. L. Bessemer.-Abbott, C. E. Dobbs, G. G. Ferguson, V. Salmon, H. S. Schaber, C. F. Birmingham.-Adams, J. H. Aldrich, T. H. Aldrich, T. H., Jr. Allen, A. W. Ball, E. M. Bonnyman, J. Bowron, J. Bush, M. W. Crawford, G. G. Crockard, F. H.

Jan 1, 1917

By Leland Wemple

PREVIOUS to 1915, zinc refining had not become a general practice among the zinc smelters in the United States. Such refining as had been carried on was confined chiefly to remelting very high-leaded zinc, such as third-draw metal from the retort furnaces, and slab dross, and hard spelter from galvanizing pots. In this remelting operation, the lead and iron were settled out and the refined spelter obtained graded fifth class, i.e., Prime Western. Owing to the presence of lead in most zinc ores, the bulk of the spelter produced by direct smelting is of this Prime Western grade. Careful selection of ores of low lead content gave purer grades of spelter, i.e., Brass Special and Intermediate; and smelters which could secure ores carrying practically no lead produced high-grade spelter. Thus, in general, the grade of spelter produced by any smelter was pre-determined by the purity of the ore it could obtain. The needs of the warring nations of Europe for enormous quantities of cartridge and other high-grade brass resulted in such a demand for the purer grades of spelter that the available supply of lead-free and low-lead ores could not produce the quantity required. Prime Western grades went begging, so to speak, while high grades sold at times from 5 to 15 c. per pound over the Prime Western market. The great demand, coupled with the wide market margin, which permitted the inherent costs and losses, resulted in the conversion of the lower grades into higher grades by processes of refining. These conditions also accentuated the development and size of the electrolytic plants, which were being worked out at the time this increased demand for pure grade of metal occurred. The electrolytic process, where available, produces a higher grade of spelter from impure ores than does the fire process; a large portion of those impurities in the ore, which by the distillation process would be carried into the spelter, are removed in the leaching of the ores and the purification of the solutions. The refining process, adopted at a number of smelters for the purpose of converting low-grade spelter into "grades suitable for brash was that of redistillation.

Jan 11, 1917

By W. H. Gauvin, N. J. Themelis

Jan 1, 1963

By H. E. Lee, P. C. Feddersen

Greenockite is the only known cadmium mineral of importance. It occurs rather universally, in minor concentrations, as a secondary mineral in sphalerite deposits. The world's cadmium output is obtained through the processing of metallurgical by-products, largely from the treatment of residues from electrolytic zinc, retort zinc and lithopone plants. These sources are supplemented by the processing of fumes from lead and copper smelting operations. The development of modern selective flotation practice in the decade 1920-1930, which permitted the economical mining of complex lead-zinc ores, resulted in significant increases in the quantities of cadmium entering lead smelting systems. Being closely related to zinc as to occurrence, properties and production, most detailed description of cadmium recovery methods are to be found recorded in connection with zinc metallurgy. Other than occasional articles pertaining to particular operational procedures, literature offers but little in the nature of a balanced survey of lead smelter cadmium recovery practices. While many of the basic operations described for the recovery of cadmium from zinc by-products are applicable to the treatment of lead plant products, the inherent problems involved differ widely. In general the cadmium content of related by-products from routine lead smelting operations is present in lower concentrations, exists in a less soluble state and is associated with both a greater quantity and a greater variety of detrimental impurities. To cope with these problems, lead smelter practices are found to follow the general outline: Preparatory Processing 1. Concentration operations 2. Sulphation operations Cadmium Plant Processing 1. Leaching operations 2. Purification operations 3. Sponge precipitation operations 4. Metal recovery operations 5. Refining and casting operations As in the case of related zinc plant operations, cadmium recovery practices at lead smelters are not standardized. They not only vary as to type, but also extent. Depending upon prevailing conditions, lead smelter cadmium operations range from simple concentration campaigns, for the purpose of sufficiently "up-grading" products for shipment elsewhere, to complete processing steps for the production of refined metal. Preparatory Processing The cadmium content of lead smelter receipts is low and, as a rule, proportionate to the zinc content; the usual range of cadmium contained being of the order of 0.01-0.05 pct. Were it not for the low boiling point of cadmium, such small concentrations would, no doubt, be lost in the large tonnages of slag, metal and other smelter end products. However, the ready volatility of cadmium and its compounds at prevailing lead smelting temperatures results in its concentration in fractional portions of fume collected. This collected fume comprises a circulating load within the smelter system. Thus, the cadmium content of blast furnace fume* increases with each successive circulation until an equilibrium value is reached when the sum of the cadmium losses, due to handling and in slag, waste gases and other end products, becomes equal to the intake as ore. With ore receipts averaging, say 0.03 pct cadmium, the concentration value obtainable, through fume circulation in a routine manner, approaches 10-12 pct. In such operations, cadmium concentrations in blast furnace fume of from 3-5 pct are readily .attainable. However, the concentration gain beyond this range, with each additional circulation, is progressively decreased as a result of mounting losses occurring through handling and in end products. Therefore, to avoid excessive cadmium loss and to enhance the ultimate concentration anttainable, it is customary practice to isolate blast furnace fume at some intermediate cadmium content for special concentration procedure. The most common type cadmium concentration "campaign" involves special smelting operations wherein a relatively high portion of blast furnace fume at 3—6 pct cadmium is incorporated into the sinter charge. This type practice is roughly illustrated by the diagram on p. 111. Cadmium fume collected from lead blast furnace operations is not amen-

Jan 1, 1950

By Jed H. Mosgrove, Paul M. Budzak

INTRODUCTION Good mine ventilation starts at the drawing board. A coal mine will either be adequately ventilated or not, depending upon how good a job was done in the planning. Poor planning will not only result in inadequate ventilation, but may also lead to low productivity. A mine foreman cannot achieve high productivity while trying to maintain 4.2 m3/s (9000 cfm) of air at his last open crosscut. PLANNING What will provide adequate ventilation? A rule of thumb in laying out a coal mine is to provide enough entries in the main entry system to keep air velocities at 2.5 m/s (500 fpm) or less; at this rate of flow, good ventilation may be expected. The intake and return air courses of a coal mine do not improve with age. Entries will deteriorate, ribs will slough, roof will fall, and clutter will accumulate. The resistance offered to air travel will continue to increase with each passing year. Providing enough airways for the present and the future cannot be emphasized too strongly. In order to better appreciate good planning and bring it into better focus, assume a hypothetical mine has the following features: entries 1.5 m (5 ft) high and 6.1 m (20 ft) wide; one intake and one return entry; the mine development is 4.8 km (3 miles) from the intake and upcast ventilation shafts; and the ventilation requirement is 141.6 m3/s (300,000 cfm). Air travel of 4.8 km (3 miles) into a mine and the same distance back to the fan is not unusual. However, to do so with one intake and one

Jan 1, 1981

By R. B. Adamson

The fatigue behauior of irradiated copper single crystals is compared to that of unirradiated crystals. Proton or electron irradiation did not substantially change the fatigue-lzj.e, fatigue-hardening-, or slip-line characteristics, while neutron irradiation did appear to alter these parameters. Postfatigue transrrzission electron microscopy did not reveal any marked structural differazces between irradiated and unirradiated specimens. Possible dislocatio~a mechanisnzs to explain the observed behavior are discussed. A LTHOUGH considerable work has been done to ascertain the effects of radiation on the tensile properties of metals (e.g., Refs. 1 and 2), little work of a basic nature has been reported on the effect of radiation on fatigue properties. Since it is known that fatigue properties are sensitive to conditions at the surface of the metal, it was felt that charged particle irradiation, whose penetration into metals is very small, might influence the fatigue behavior. To investigate this surface effect, irradiations with 10 mev protons and 1.5 mev electrons were conducted. The range of 10 mev protons in copper is about 0.236 mm; which is very small compared with the thickness of the specimens used (1.53 mm). The range of 1.5 mev electrons is somewhat larger, on the order of 0.665 mm.4 On the other hand, irradiation with reactor neutrons produces damage throughout the entire specimen, so neutron irradiation can be used to study the effects of bulk damage. The damage introduced by proton and neutron irradiation is expected to con!ist of large interstitial dislocation loops,5 tiny (<25A) vacancy loops," and/or various vacancy-rich regions called depleted zones.&apos; In the case of electron irradiation at about -173°C the damage is expected to consist of isolated Frenkel pairs.&apos; The effect of radiation damage on the fatigue properties of some engineering materials has been reported,g,10 but the results are not clear-cut or consistent. SPECIMEN PREPARATION Zone-refined copper was used to grow single crystals by a modified Bridgeman technique in an inert helium atmosphere. Using a seeding method, single crystals were obtained as bars about 6 in. long with a cross section of 3 by $ in. Immediately upon removal from the split-mold graphite boat in which they were grown, the very soft bars were embedded in a room-temperature setting plastic. They could then be cut into specimen lengths without introducing any long-range damage. Generally, four or five specimens plus one seed were obtained from each bar. The plastic material was dissolved away in acetone and the orientation of each sample was checked by the standard Laue X-ray

Jan 1, 1968

By Thomas T. Read

The purpose of the following research, as originally planned, was to investigate the influence of temperature upon the plate-amalgamation process. In order to consider the amalgamation process intelligently, it was first necessary to learn the nature of an amalgam. In the performance of this task it was found necessary to consult a large volume of literature and to perform experimental investigations. The conceptions of the nature of amalgams thus obtained have so important a bearing on the amalgamation-process as a whole, as well as upon the possible influences of temperature, that it was advisable to include them in the treatment of the subject. The broader title is therefore used, even though some features of the amalgamation process have been treated very briefly and others entirely omitted, there being nothing new to present regarding them and their development not required for a clear presentation of the matter in hand. I. Historical Introduction. Gold was one of the earliest metals known to man. Occurring in the metallic state, it could be picked up in nuggets or mashed, as gold dust, from the sand and gravel of streams. Interestiag descriptions of the processes employed in early times for the recovery of gold may be found in Pliny&apos;s Natural History and in the treatise by Agricola on ores and minerals. But little progress in the treatment of the ores of gold mas possible until efficient crushing machines were devised. This want was supplied by the introduction, about the beginning of the sixteenth century, of the stamp-mill to treat the ores of Saxony. Blanket-strakes, or similar devices, were used to catch the

Jan 1, 1907

By V. H. Wilhelm

Developments in California during 1942 were marked by many difficulties in operation, of which the lack of labor and material were the main factors in slowing down work. During the many years of curtailment, which has been effective since 1928, a number of estimates of the state&apos;s capacity to produce have been built up, but during the year 44 fields were &apos;eliminated from curtailment and valuable data were obtained as to the normal capacity of these fields to produce. Because of new discoveries and the normal decline of old fields, which is more than 15 per cent per year, it is estimated that the sustained capacity of the state for a 5-year period is not more than 500,000 bbl. per day. Decline in the rate of production was greatest in the fields in the Los Angeles Basin district. These fields have been produced nearly at capacity for many years and had no reserve of shut-in production. The San Joaquin Valley and Coastal districts have responded well to the increased allotments assigned to them by the Petroleum Administrator for War. All heavy oil fields in the state are on full production, except in the Coastal district, where transportation is the limiting factor. Although approximately 2500 more wells are on production in the state than there were in 1941, the supply of heavy crude has not yet reached the demand. A summary of statistics for 1941 and 1942 is given in the following table. Production in California for 1942 was 680,000 bbl. a day as against 631,000 bbl. a day for 1941, an increase of 7.8 per cent. The present maximum efficient rate has been estimated at 808,000 bbl. per day. This indicates that in the 15 restricted fields of the state there is only an additional 83,000 bbl. per day of light oil available. Drilling Activity Drilling activity during the year was equivalent to but 70 per cent of normal. Although shortages of equipment and labor were in part responsible for the reduced drilling activity, the main reasons were the lack of proved locations due to a cycle of very few new discoveries and the restrictions of Conservation Order M-68. For the first six months of the year, 300 new oil wells were completed, which were equally divided as to the number of light and heavy oil producers. But, for the last six months of the year, it is

Jan 1, 1943

By V. H. Wilhelm

Developments in California during 1942 were marked by many difficulties in operation, of which the lack of labor and material were the main factors in slowing down work. During the many years of curtailment, which has been effective since 1928, a number of estimates of the state&apos;s capacity to produce have been built up, but during the year 44 fields were &apos;eliminated from curtailment and valuable data were obtained as to the normal capacity of these fields to produce. Because of new discoveries and the normal decline of old fields, which is more than 15 per cent per year, it is estimated that the sustained capacity of the state for a 5-year period is not more than 500,000 bbl. per day. Decline in the rate of production was greatest in the fields in the Los Angeles Basin district. These fields have been produced nearly at capacity for many years and had no reserve of shut-in production. The San Joaquin Valley and Coastal districts have responded well to the increased allotments assigned to them by the Petroleum Administrator for War. All heavy oil fields in the state are on full production, except in the Coastal district, where transportation is the limiting factor. Although approximately 2500 more wells are on production in the state than there were in 1941, the supply of heavy crude has not yet reached the demand. A summary of statistics for 1941 and 1942 is given in the following table. Production in California for 1942 was 680,000 bbl. a day as against 631,000 bbl. a day for 1941, an increase of 7.8 per cent. The present maximum efficient rate has been estimated at 808,000 bbl. per day. This indicates that in the 15 restricted fields of the state there is only an additional 83,000 bbl. per day of light oil available. Drilling Activity Drilling activity during the year was equivalent to but 70 per cent of normal. Although shortages of equipment and labor were in part responsible for the reduced drilling activity, the main reasons were the lack of proved locations due to a cycle of very few new discoveries and the restrictions of Conservation Order M-68. For the first six months of the year, 300 new oil wells were completed, which were equally divided as to the number of light and heavy oil producers. But, for the last six months of the year, it is

Jan 1, 1943

By H. K. Najarian

This paper describes recent development work on the Wheaton-Najarian Vacuum Condenser for Zinc at Josephtown Zinc Smelter, wherein, to dissipate the heat of condensation, the "airlift" action of gases bubbling through the molten zinc in the condenser is utilized to circulate the zinc through an outside "cooling well" and cooling it continuously. AN earlier paper1 gave an account of initial development of the Wheaton-Najarian Vacuum Zinc Condenser at the Josephtown Zinc Smelter of the author&apos;s company. This paper is a progress report of further development work over the years directed toward improving the efficiency and increasing the capacity of the condenser. In this type of zinc condenser, zinc vapor with accompanying noncondensable gases from electro-thermic zinc smelting furnaces is passed upwardly through a mass of molten zinc confined in a refractory-lined U-tube. The zinc vapor condenses in the molten zinc and the noncondensable gases are drawn out of the condenser by suction applied over the surface of the molten zinc. The heat of condensation of zinc, amounting roughly to 1000 Btu per lb of zinc condensed, is absorbed by the molten zinc and, in turn, is dissipated continuously to keep the temperature of molten zinc in a range to permit effective condensation. At the time of publication of our first paper, we had at Josephtown Smelter five units of 69-in.-bore metal furnaces and one larger metal furnace having 96-in. bore. Each 69-in. metal furnace had a condenser with 22-ft overall length. The 96-in.-bore furnace had a condenser with overall length of 31 ft. Condensers were of circular cross section, with corrugated steel outer shell and carborundum lining. Diameter of these condensers was 32 in. inside the refractory lining. Molten metal in the con- denser was kept cool by spraying cold water on the outside of the condenser shell. Under good operating conditions, 22-ft condensers produced some 12 tons of metal per day; and the 31-ft condenser, about 22 tons. When the furnaces produced zinc vapor at higher rates, invariably the condenser temperature would rise to levels where excessive blue powder would form and accumulation of accretions in the condenser would be accelerated. We found that, with the system of cooling in use, we had to provide heat-exchange surface of some 150 sq ft to condense 1000 lb of metal per hr with reasonable efficiency. The wall of the condensers was made up of 5/8-in.-thick corrugated-steel shell having next to it several layers of mica and inner refractory lining of special carborundum shapes to fit the corrugation in the steel shell, and having a minimum thickness of about 2 in. Relatively low rate of heat transfer through the thick condenser wall and slow internal circulation of the molten zinc limited the rate of heat exchange between hot metal within the condenser and the cold water flowing over the outer steel shell. The thickness of the condenser wall was about minimum that we considered suitable for continuous operation. As we learned to operate the electrothermic furnaces with higher power input, producing zinc vapor at higher rates, it became apparent that we needed condensers having larger heat-exchange surfaces. Our 96-in.-bore furnace had 6 single-phase power circuits with transformers rated at 500 kva each and normally good for power input of 3000 kw.

Jan 1, 1959

By J. R. Lewis, J. H. Hamilton, V. D. Hanson, K. W. Heinlein, J. M.194-000-000-006 Vonfeld, J. P. Denny

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion NO. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and — -in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3h-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 34 in., and 1&apos;/4x in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into YSXO which is cleaned in the fine coal plant and 3x for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1953

By J. H. Hamilton, J. R. Lewis, K. W. Heinlein, V. D. Hanson, J. M. 194-000-000-006 Vonfeld, J. P. Denny

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion NO. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and — -in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3h-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 34 in., and 1&apos;/4x in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into YSXO which is cleaned in the fine coal plant and 3x for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1953

LIST OF THE MEETINGS OF THE INSTITUTE AND THEIR LOCALITIES FROM ITS ORGANIZATION Transactions Number Place Date Vol Page 1 Wilkes-Barre, Pa May, ?71 1 3 2 Bethlehem, Pa August, ?71 1 10 3 Troy, N Y November, ?71 1 13 4 Philadelphia, Pa February, ?72 1 17 5 New York, N Y May, ?72 1 20 6 Pittsburg, Pa October, ?72 1 25 7 Boston, Mass February, ?73 1 28 8 Philadelphia, Pa May, ?73 2 3 9 Easton, Pa October, ?73 2 7 10 New York, N Y February, ?74 211 11 St Louis, Mo May, ?74 3 3 12 Hazleton, Pa October, ?74 3 8 13 New Haven, Conn February, ?75 3 15 14 Dover, N J May, ?75 4 3 15 Cleveland, 0 October, ?75 4 9 16 Washington, D C February, ?76 4 18 17 Philadelphia, Pat June, ?76, 5 3 18 Philadelphia, Pa October, ?76 5 9 19 New York, N Y February, ?77 5 27 20 Wilkes-Barre, Pa May, ?77 6 3 21 Amenia, N Y October, ?77 6 10 22 Philadelphia, Pa. May, ?78 6 18 23 Chattanooga, Tenn May ?78 7 3 24 Lake George, N.Y. Oct. ?78 7

Jan 1, 1923

LIST OF THE MEETINGS OF THE INSTITUTE AND THEIR LOCALITIES FROM ITS ORGANIZATION Transactions Number Place Date Vol Page 1 Wilkes-Barre, Pa May, ?71 1 3 2 Bethlehem, Pa August, ?71 1 10 3 Troy, N Y November, ?71 1 13 4 Philadelphia, Pa February, ?72 1 17 5 New York, N Y May, ?72 1 20 6 Pittsburg, Pa October, ?72 1 25 7 Boston, Mass February, ?73 1 28 8 Philadelphia, Pa May, ?73 2 3 9 Easton, Pa October, ?73 2 7 10 New York, N Y February, ?74 211 11 St Louis, Mo May, ?74 3 3 12 Hazleton, Pa October, ?74 3 8 13 New Haven, Conn February, ?75 3 15 14 Dover, N J May, ?75 4 3 15 Cleveland, 0 October, ?75 4 9 16 Washington, D C February, ?76 4 18 17 Philadelphia, Pat June, ?76, 5 3 18 Philadelphia, Pa October, ?76 5 9 19 New York, N Y February, ?77 5 27 20 Wilkes-Barre, Pa May, ?77 6 3 21 Amenia, N Y October, ?77 6 10 22 Philadelphia, Pa. May, ?78 6 18 23 Chattanooga, Tenn May ?78 7 3 24 Lake George, N.Y. Oct. ?78 7 25 Baltimore, Md. Feb. ?79 7

Jan 1, 1923

By C. A. Stetefeldt

The purpose of this paper is to convince the managers of silvermills that their customary reports in regard to the percentage of silver saved are far from the truth. It may be unpleasant for them to have the proofs of this fact brought before the public, and for this rezson the author refrains from giving the names of the mines and mills from which the statistics here recorded were obtained. It may be asked: What is gained by informing the stockholders of our silver-mines that they receive incorrect reports? Will this increase their dividends or lessen their assessments? In the first place, it is always better to publish honest facts than fiction. In the second place, many superintendents believe in the correctness of their reports, and have no idea how much they deceive themselves. Finally, the knowledge that hitherto unexpected losses of silver take place may lead to the adoption of methods by which these losses are greatly reduced. These remarks refer also to smelting-works. The present paper, however, is concerned exclusively with amalgamation and lixiviation. The first and most important error is found in the inaccuracy of the commercial assay; the second in the incorrect sampling of tailings in amalgamation. Further, mills not doing custom-work, but reducing exclusively ores from mines to which the mills belong, are in the habit of weighing the ore only approximately and without making moisture-tests. Thus, a correct basis for calculating the actual percentage of silver extracted is lacking. The superintendent, then, finds it convenient to report the apparent extraction, i.e., the extraction indicated by comparing the value of the roasted ore, minus soluble salts, with the value of the tailings. How many errors are thereby committed it is not necessary to point out.

Jan 1, 1895

By D. H. Fleming

DURING 1950 the jet-piercing process was used commercially in the piercing of primary blast- holes in magnetic taconite at the preliminary taconite plant of the Erie Mining Co., Aurora, Minn. The Erie Mining Co. purchased a jet-piercing ma- chine and associated equipment in 1949, and the machine has been in operation since its installation in the summer of that year. An overall view of the pit with the JPM-1 jet-piercing machine in operation is shown in Fig. 1.

Jan 7, 1951

By Nathaniel Herz

The use of zinc dust for precipitating the pecious metals from cyanide solutions is well established now in many places, and has many advantages over the shavings method of precipitation. Although much work has been done o,n methods of testing zinc dust, as yet. no absolutely positive test has been devised to determine the value of the material. Certain facts are known, but often it happens that results on a working scale are not as expected from the laboratory tests. The purpose of this paper is to record some obscrvations made while testing a large variety of zinc-dust samples in the laboratory. W. J. Sharwoodl has described standard analytical methods, which have been followed, with the changes noted. The regular determinations include zinc oxide, by solution in ammonia and ammonium chloride, lead, and precipitating efficiency. In some cases cadmium was determined by a method similar to that described by J E Clennell.2 I will discuss the precipitating-efficiency test in detail. The procedure is as follows: "A solution of potassium silver cyanide is prepared by dissolving 10 grams of silver cyanide (AgCN) and 5 grams of 99 %&apos; potassium cyanide in a little water and diluting up to 1000 cc. It is then adjusted by addition of a little more XCN or AgCN until the solution indicates from 0.12% to 0.15% free KCN by titration with standard silver nitrate. The titration is best made by using a 10-cc. or 20-cc. sample, adding 1 CC. of 2 % potassium iodide, and a slight excess of ammonia, as the end point is then sharper. Or 15 grams of pure crystallized XAg(CN)2 may be dissolved in a liter of water and 1.5 grams XCN added. " Weigh out 0.5 gram zinc dust into a 300-cc. beaker. Add a few cubic centimeters of watcr and stir until zinc is well mixed, then pour in 250 cc. of the prepared solution, stirring vigorously. See that all lumps are broken up, and continue stirring for fully five minutes. Stir occasionally (at least every 10 minutes) until the end of two hours from the addition of the solution. Then filter upon an 11-cm. filter, wash precipitate thoroughly, sprinkle with test lead, wrap it carefully in the paper, place in a scarifier with about 20 grams test lead, burn paper cautiously in muffle, scorify five minutes, cupel at low temperature and weigh silver. Milligrams silver obtained from 0.5 gram zinc X 0.0606 = percentage precipitating efflciency."³

Jan 1, 1916

By A. T. Janssen

Production planning of direct-shipping ore from a number of relatively small deposits serves to sequence development and mining of these deposits so that both long- and short-range demand can be met while the timing of development cost is such that the average annual rate is minimized. Hand-computed methods have the disadvantage that they require substantial effort and time and cannot be exhaustive. The application of operations research and computers not only overcomes these disadvantages, but opens up additional applications, such as feasibility studies involving alternative mining strategies and changes in the ore market. The method employed concerns linear programming and specifically the Big-M algorithm applied to the variant on the simplex technique with penalties applied to deviations from the desired ore grade. Acceptable 20-year production schedules could be generated in relatively little time, i. e. , in less than 200 seconds for the central processor of the Control Data Corporation 6400 system of the University of Colorado in Boulder.

Jan 1, 1969