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By R. S. Shoemaker, R. M. Darrah

Expanded markets for copper in the past few years and a consequent search for new ore bodies have revitalized the widely known but seldom applied method of producing copper called heap leaching. Heap leaching is defined here as the process applied to oxide ores which have been mined solely for the purpose of leaching. In this case, all costs of mining the ore become a part (and a substantial one, as will be shown) of the cost of the copper produced, whereas in dump leaching the mining cost has already been accounted for as a stripping cost for the sulfide ore. Of course, with newly opened ore bodies, stripping costs may be divided through accounting procedures, between the copper produced by the mill and that obtained by leaching, but in general this type of operation must be economic on the basis of the sulfide ore alone. Heap Leaching Requires Low Capitalization Heap leaching, as a beneficiation method, is unique in that it is the only process used for the recovery of a metal in which the time involved from mining the ore to metal production is measured in months instead of hours or at the most, days. Then too, after the ore body has been depleted and mining has ceased, metal production will continue for some substantial period of time. Heap leaching is also unique because unlike any other method of beneficiation, metal recovery cannot be accurately forecast ahead of time and recovery will never approach that of a conventionally milled sulfide ore or even that of a vat leached oxide ore. Counter-balancing this drawback, however, is the relatively low capital expenditure necessary to put a heap leaching operation into production and the consequence that smaller ore bodies which would not support the higher capital cost involved in a vat leaching plant can be made into producers. Stated more simply, in heap leaching the operator accepts low recovery in exchange for low capital costs and low break even point. Whether the operator chooses vat leaching with its higher recovery and higher capital costs, is a mater of economics which depends in part on the nature of the ore to decide. For instance, an ore that will vat leach in 7 to 10 days using 7 pounds of sulfuric acid per pound of copper produced might use double or triple that amount of acid during the months that are required for heap leaching.

Jan 12, 1968

By W. S. Row

This mine is operated at 4000 to 4500 tons daily through a single shaft, with one rock hoist and one service hoist. Latest shaft construction is concrete with wooden dividers. Economics of drifters and air-leg drills are compared, as well as costs of shrinkage and sub-level stoping. Economy of large haulage equipment and underground crushing is shown.

Jan 3, 1953

By Edward L. Fidler, Thomas E. Finch

Thin seam coal mining in the western US nominally equates to recovery of stray seams associated with thicker, major seams. The thin or stray seams encountered are generally recovered, but not in all cases. The alternative to extraction is, of course, spoiling. The majority of thin seams are encountered somewhere in the overburden above the major seam or in the parting between thick seams. Although some are true strays and appear anywhere in the geologic sequence, others might be classified as thin "rider" seams immediately above or below a thick seam and separated by some equally thin parting. For the purpose of this discussion, stray or thin seams are defined as those which are less than 1.2 m thick or represent less than 10% of the expected extractable coal. [Figure 1] is a graphic representation of typical thin, stray, or rider seams which actually appear in a western operation. Operationally, stray seams create irritating scheduling problems and require special handling techniques which do not always lead to efficient equipment utilization. Additionally, coal recovery from a thin seam is considerably less than the 95% usually recovered from thick seams. Higher recovery in thin seams leads to dilution problems, which depreciates Btu quality, and most western coals are not in a position to lose heating value. Thin western seams also undulate and pinch or swell more than thick seams, and this leads to more difficult recovery. Often they are associated with equally dark carbonaceous shales and visual selectivity is difficult at best during daylight hours. Resource Recovery Regulations Thin seam recovery in western mines is encouraged by both federal and state regulations and by the mining industry's concern for total resource recovery. [Table 1] lists the western coal states and their recovery regulations. The federal regulations are found in the Surface Mining Control and Reclamation Act of 1977. They state, "General performance standards shall . . . require the operation as a minimum to … conduct the surface coal mining operations so as to maximize the utilization and conservation of the solid fuel resource being recovered." Economic stimulation for thin seam recovery appears to be marginal or at the least controversial. Scheduling and operational problems often outweigh the nebulous profits created by engineering analysis. The only alternative to extracting thin seams occurring above major production seams is to strip directly through them and lose the resource to the spoil pile. Although this is not palatable to environmental emotions or resource promoters, it is the most simple expedient to operations people.

Jan 1, 1981

By William H. Waggaman

CURTAILMENT of the mineral industry as a whole undoubtedly will follow world peace, but the output of certain minerals should pursue a course well above the average on any curve of probable output projected beyond the trends exhibited through 1939. Without speculating on industrial conditions in the pending recomversion period, demand for fluorspar should be greater than before the war because of these factors, introduced by or resulting directly from military necessity: (1) Discovery and development of new deposits. (2) Improved technology in benefication. (3) New and expanded uses for fluorspar and its derivatives. The steel industry has been and continues to be the greatest consumer of fluorspar, and it wag largely to meet the wartime needs for steel that the domestic fluorspar industry has been expanded so greatly. Not only were the search for and development of new deposits of this mineral considered esstinal but the marginal nature of some of the,c- domestic deposits made it necessary to improve mineral dressinc and processing techniques to utilize them

Jan 1, 1945

By A. S. Richardson

THE design of mine airways receives, in general, very little engineering treatment. To a large extent this is, of course, due to the fact that information upon which to base calculations is seldom available in definite form. In many mines the extent of the workable deposits may not be known at the time an air course is planned, and the length of service is indeterminable. Volume of air required for ventilation may also increase with extension of the life of a property, and natural obstacles, such as heavy ground, often place limitations upon the size of opening that can be maintained. Perhaps a majority of the shafts used in ventilation of metal mines were originally intended as hoisting shafts, and many of the tunnels and cross-cuts similarly used, were also designed as haulage ways or for other operating purposes. Even when a shaft is sunk solely for ventilation, custom usually dictates a size and shape consistent with the type commonly used for ordinary operations, because experience has demonstrated it can be excavated and maintained. In fact it may not be very far wrong to say most ventilation openings were never made with any consideration as to economic design for such use. When an air course is required to pass air between different parts of mine workings under a fixed pressure, the size of opening required for a given volume of flow, is, of course, determined by the resistance factor of the air course. It is then impossible to proportion the opening with reference to economics of power consumption, and this paper is limited to the design of openings where ventilation pressure or power consumption may be varied for economic advantage.

Jan 2, 1926

By E. Klar, A. B. Michael

Differences in the electrical conductivities of copper powder sintered under reducing, selectively oxidizing, and neutral atmospheres are related to impurities in solution or as precipitated oxides. The precipitation of impurities as oxides during sintering in nitrogen is proposed for maximizing the conductivity of sintered copper. Conductivity equations for two-phase systems are summarized. Selected equations are applied to porous sintered copper and composite structures. A recent review of the influence of impurities on the electrical conductivity of copper by Gregory et al.1 emphasized that an impurity in solid solution has a much more pronounced effect on reducing the electrical conductivity than when present partly or wholly as a second phase. When impurities in solid solution can be precipitated as oxides, the copper is purified with respect to these elements and the conductivity is increased. Cast and wrought copper, therefore, frequently contain an intentional residual oxygen concentration to oxidize and precipitate impurities less noble than copper. Copper powders generally also contain impurities which can contribute to a reduction in the electrical conductivity of the sintered material. The most deleterious impurities commonly found in commercial copper powders which markedly decrease the electrical conductivity when in solid solution but which can be precipitated as oxides include iron, tin, antimony, arsenic, cobalt, and nickel. In addition to impurities, porosity in sintered copper also contributes to a reduction in the electrical conductivity. The work reported herein discusses the influence of impurities, sintering atmospheres, and porosity on the electrical conductivity of sintered copper. These are important factors for controlling the electrical conductivity of materials such as sintered copper electrical contacts. Several publications on the electrical conductivity of sintered copper and composite materials with random pores or obstacles have incorrectly considered the conductivity to be proportional to the volume fraction of conducting material. However, analyses and equations have been proposed, the earliest of which is perhaps Lord Raleigh's of 1892,' which more accurately describe the conductivity of two-phase systems. These equations which in many cases allow one to closely estimate the electrical conductivity of porous sintered materials and two-phase composites will be reviewed and related to the measured electrical conductivity of sintered copper, copper-graphite. and Ag-W composites. INFLUENCE OF IMPURITIES AND OXIDIZING, REDUCING, AND NEUTRAL ATMOSPHERES Zone-Melted and Leveled Copper. The electrical conductivities of an electrolytic copper and a lower-purity compacting grade of copper powder after consecutive treatments in reducing or selectively oxidizing atmospheres are compared in Fig. 1. The powder and drillings from the electrolytic copper ingot were melted and solidified in a graphite crucible under hydrogen. The vertical zone leveling technique of multiple passes in opposite directions was used to obtain a uniform distribution of impurities. The electrical conductivity was measured on machined specimens 3 by 0.10 in. diam of the zone-leveled material and after the same specimens were consecutively treated as follows: 1) heating in hydrogen at 1600°F for 3 hr; 2) heating in air in a sealed tube so that 0.04 pct 0 was introduced; heating was started at 1000°F for 1 hr and continued at 1800°F for 8 hr; the stepwise diffusion treatment was used to avoid loss of copper due to evaporation of copper oxide at higher temperatures; and 3) final heating in hydrogen at 1600°F for 3 hr. A L&N Kelvin Bridge, type 4306, and a test fixture with knife edges 2 in. apart were used to measure the room-temperature conductivity with an estimated accuracy of ±0.5 pct. In all cases the conductivity of the electrolytic copper was approximately 100 pct of IACS. The conductivity of the impure copper was 80 pct of IACS both in the cast state and after heating in hydrogen. However, after the heating in air, the conductivity of the impure material was 100 pct IACS. After again heating in hydrogen, the electrical conductivity decreased to 60 pct of IACS. This decrease is attributed both to the dissolution of impurities and observed intergranular cracks due to the phenomenon of hydrogen embrittle-ment in copper. These data show that the electrical conductivity of commercial copper as represented by a compacting type powder can be increased significantly by the precipitation of impurities as oxides through heat treatment in an oxidizing atmosphere. Sintered Copper Powder. A commercial compacting type of copper powder pressed to various densities was sintered either in nitrogen, dissociated ammonia, or dissociated ammonia followed by a selectively oxidizing atmosphere of air in sealed Vycor tubes so that 0.06 pct O was added to the material. The electrical conductivity was measured perpendicular to the pressing direction on as-sintered specimens of approximately 3 by 0.25 by 0.15 in. The fully dense material was obtained by zone melting and leveling

Jan 1, 1969

By Robert L. Grayson, Y. J. Wang, R. Mike Mishra

Prediction of ventilation needs in long range mining projections requires serious consideration when formulating those projections. Predicting the behavior of the ventilation system of an old, complex mine poses an especially provocative problem. Air-flow around and through mined-out areas via bleeder systems, leakage flows through stoppings or other air control devices, and the use of multiple fans all add to the basic complexity of a ventilation system and make predictions of ventilation changes very difficult. Fortunately, the task of making ventilation decisions can be greatly assisted by ventilation surveying and computer simulation. The Nemacolin Mine of Jones and Laughlin Steel Corporation, located in Southwestern Pennsylvania, is one such old mine which used the computerized approach successfully. A mine-wide pressure-quantity survey was used to obtain realistic values for airway resistance factors. These values, together with network topology, formed the basic input for computer simulation. Results from this simulation indicated that air quantities generated at critical airways by the model were generally within a couple of percent of those measured in the actual survey. It was then concluded that the model was acceptable for practical application, and it was used immediately for studying the effects on the ventilation system of changing the blade settings of fans. Additionally, the model was modified to incorporate projected mine development and used to predict future ventilation.

Jan 1, 1982

John S. Kennedy, Chamhersburg, Pa. (communication to the Secretary): The washing-plant described by Mr. Johnson is a good illustration of the advantages derived from a well-designed and care-

Jan 1, 1895

By Charles Janin

The first successful bucket-elevator dredge to operate in California was put in comnlisvion wt Oroville in March, 1898. There had been numerous previous attempts at dredging, but noue of the earlier boats proved a success. The gold-miners in California earl3 conceived the idea of a machine to dig gravel from the beds and bars of auriferous streams that were inaccessible by the methods then employed, and it was only a few months after the discovery of gold in California that such a machine was Shipped around Cape Horn from New York to San Francisco. Thij was, however, but the forerunner of many failures in gold-dredging, and was soon at the bottoni of the Sacramento river. During succeeding years many other unsuccessful attempts mere made, and it was not until 1897 that a dredge of the single-lift bucket-elevator type mas floated on the Yuba river. This dredge was built by the Risdon Iron Works for 12. H. Postlethwaite, and would probably have been a success if it had been operated on some of the rich Ororille ground

Jan 1, 1912

By F. F. Poland

The purpose of this paper is to describe the recent applications and improvements made in the process and equipment for the recovery of metallic zinc from secondary metals by means of high temperature electric resistor furnace distillation. The process and high temperature furnace, as developed by Revere Copper and Brass Incorporated, Research and Development Department, was described in a previous paper. † Since the publication of that paper the process and equipment have been licensed to others. Consequently its field of application has been expanded, particularly for the treatment of scrap brass, galvanizers dross and scrap zinc. Other applications are being considered, particularly the melting of metallic titanium and special alloys requiring a controlled atmosphere. Consideration is also being given to modification in design that would allow the melting of copper cathodes for the production of oxygen-free and electrolytic tough pitch copper shapes without the necessity of blowing and poling. In the design of a furnace for melting copper cathodes it is intended that the fundamental principle of charging the solid material into a large pool of molten metal through a double door charging vestibule as shown in Fig 3 of the previous paper, † would be adhered to and, in addition, either dehydrated and desulphurized producer gas or commercial nitrogen, that is, approximately 96 pct nitrogen, 2 pct hydrogen 2 pct CO, also dehydrated, would be used to provide a protective atmosphere for the resistors and to flush out more soluble gases from the molten copper. The operation of a small unit producing approximately 1000 lb of copper per hr has shown that the commercial nitrogen will allow the production of tough pitch electrolytic copper when the metal is transferred from the meltdown unit contemplated to a special type induction unit where the oxygen absorption and temperature can be controlled. Standard practice in treating low grade scrap brass has been to oxidize the zinc by blowing air through the molten metal. This operation results in the recovery of approximately two-thirds or less of the zinc content of the brass in the form of an impure zinc oxide having only a fraction of the value of the zinc in metallic form. The balance of the zinc is lost to slag, fume and drosses. Fig 1 shows the revised flow sheet of the process for the treatment of scrap brass. For this application and as shown on the flow sheet a low frequency induction furnace has been substituted for the Wilkins-Poland type as a melting unit. This type of furnace has been found the most suitable for the purpose as will be shown by the following discussion. Serap Brass The melting of refinery scrap brass preparatory to distillation presents a melting problem quite different from the usual melting of brass mill scrap for recasting because refinery scrap is a heterogeneous mixture of alloys each having different melting and boiling points, and coated with a variety of corrosion products and foreign materials. In addition, brass and copper plated pieces of sheet iron, stainless steel, ceramic materials, and others are present that do not melt except at temperatures above the boiling point of the liquid bath. In the previous paper it is noted that difficulty was encountered in the melting of refinery grade scrap brass without excessive loss of zinc by volatilization and that 5 to 10 pct razorite flux was used to alleviate this condition. While this flux was successful in preventing the excessive loss of zinc by volatilization it interposed a low heat conductivity layer between the source of heat and liquid charge and resulted in lowering the melting rate of the furnace. This same problem presents itself in all types of furnaces in which the metal is heated by radiation and conduction from above the charge. It has been found in melting this type material that the melting rate per hour or per square foot of hearth area

Jan 1, 1950

By D. A. Dukelow, K. Li, G. C. Smith

Decarburization in the Fe-C system by oxygen top blowing has been studied in laboratory -scale experiments. It is shown that equilibrium models fail to explain or predict either the course of refining or endpoint conditions, giving results which either are incompatible with the chemistry of the system or do not satisfy material balance requirements. Also the path of decarburization was found to vary even for heats made under apparently identica1 conditions. A promising approach to analyzing the decarburization results is to relate oxygen efficiency fm carbon removal to bath carbon content. This relationship for Fe-C heats shows the same range of oxygen efficiencies as is obtained in pilot-plant and commercial heats using hot metal-scrap charges. This implies that oxygen transfer is primarily controlled by the decarburization reaction itself, independent of other refining reactions. Therefore, it should be possible to study separately decarburization and slag-metal reactions. DECARBURIZATION is probably the most important reaction in steelmaking. Not only is it a main reaction in the refining of iron to steel but it also provides the stirring action in the bath necessary for the diffusion processes to proceed at reasonable rates so as to make a steelmaking process practical. Kinetics of decarburization in the open-hearth process has been a subject of investigation for many years.&apos;-B It is generally accepted that at steelmaking temperatures the rate of homogeneous C-0 reaction is extremely high and cannot constitute a rate-controlling step. Diffusion of oxygen through a boundary film in the metal phase has been suggested by arken&apos; as rate-determining. Recently, Larsen and sordah16 concluded from experiments in a laboratory furnace that, with oxygen supplied from air or combustion gases, the rate of "steady-state" carbon boil is controlled essentially by a diffusion process of O2, Co2, or H2O through a film of nitrogen above the slag surface. Displacing this diffusion film by a stream of nearly pure oxygen produced a ten-fold increase in the rate of carbon boil with the rates of slag-metal oxygen transfer, bubble nucle-ation, and other steps all apparently able to keep pace. In the top-blown basic oxygen process, however, the transport of oxygen takes a more direct route. and the state of bath agitation is much more turbulent than in the open-hearth process. In addition, direct contact of the gas with the metal phase provides opportunity for direct oxidation of carbon. It is likely that the rate-limiting factor for the decarburization reaction will be different. However, only a few descriptive discussions of the subject have been reported in the literature.10-l2 Studies of the decarburization kinetics based on plant or pilot-plant data are necessarily complicated and are influenced by other refining reactions which occur simultaneously. In order to investigate the mechanism of decarburization, experiments have been conducted in which carbon-saturated iron melts were top-blown with pure oxygen over a range of conditions. It is hoped that this study will form a foundation on which a more basic understanding of this important reaction may be built. EXPERIMENTS One group of blowing experiments was made in a standard 200-lb induction furnace and another group in a 500-lb induction furnace. The furnaces were modified to the general shape of a basic oxygen furnace by adding a rammed refractory cone section to the regular crucible body. Crucible and cone were of high MgO (95 pct) material. A water-cooled lance, 1/2 in. in diam and threaded at one end to take a nozzle, was used for blowing oxygen. The lance with its water and oxygen lines was supported on a cantilever arrangement so that it could be moved up, down, or sideways. Oxygen of 99.5 pct purity was supplied from a cylinder and metered through a rotameter equipped with pressure and temperature gages. Another pressure gage was located at the top of the lance. A schematic diagram of the assembly is shown in Fig. 1. Before each experiment, a weighed amount of ingot iron, containing 0.02 pct C, < 0.01 pct Si, 0.10 pct Mn, 0.019 pct P, and 0.015 pct S, was charged in the furnace and melted down by induction heating. Graphite was then added to the molten charge until it became saturated. When the temperature of the charge reached the desired level, the lance was lowered to a predetermined height above the bath

Jan 1, 1964

By R. J. Brison, W. C. Bleimeister

The International Salt Company has long been interested in finding an efficient process for the removal of impurities from rock salt, and particularly from the rock salt produced at the Detroit mine. Methods which have been used previously at the Detroit mine included both hand sorting and differential crushing in a Bradford breaker with rejection of the oversize. Neither of these methods was adequately effective. The development of a new process for removal of impurities from rock salt has been described in the preceding article." Basically, the process consists of exposure of crude rock salt to radiant heating which selectively heats the impure particles. Subsequently the heated particles are removed by adhesion to a heat sensitive-coating on a conveyor belt. This method was found to be effective in improving both the quality and appearance of rock salt by the rejection of the dark, impure particles. The first commercial unit employing the new thermoadhesive method of dry separation of minerals was installed at the Detroit mine of International Salt Company in Detroit, Mich. (Fig. l). This unit was first put into operation in March 1957. The purpose of the Detroit installation was 1) to serve as a pilot plant for evaluating the process on a commercial scale and 2) to produce marketable quantities of rock salt of improved quality and ap- pearance. The design of this plant was based on the results and information obtained from a 1-tph pilot plant at Battelle Memorial Institute. DESCRIPTION OF PLANT Location: The separation plant was installed underground near the crushing and screening plant, for this was the most suitable location from the standpoint of material handling. Furthermore, ample space was availabe at this location. Because the process requires a relatively constant temperature, the uniform year-around temperature of the mine (57°F) gave an added advantage to an underground operation. Feeding: The rock salt is conveyed to a surge bin, from which it is fed to the heating unit by a conveyor. The feed-rate control consists simply of an adjustable gate at the bottom of the surge bin. The plant may be fed either No. 1 grade of rock salt (0.279 to 0.375 in.) or No. 2 grade (0.375 to 0.500 in.) with a maximum input capacity of 500 tons per 15-hour day. Heating Unit: The heating unit (Fig. 2) consists of a panel of infrared lamps suspended inside a low-speed revolving screen. The radiant-heat panel, which is curved for most effective utilization of the heat, holds 240 infrared lamps of standard industrial type, each rated at 500 watts. The position of the panel is adjustable so that it can be placed centrally over and to within 6 in. of the salt bed on the screen. The entire heat panel is suspended by rollers in ¦ R. J. Brison and O. F. Tangel: Development of a Thermo-adhesive Method for Dry Separation of Minerals.

Jan 1, 1961

By D. R. Shoemaker

BACKGROUND AND PROJECT DESCRIPTION A 16.5 kt per day underground molybdenum mine is being developed at Questa in northern New Mexico by Molycorp, Inc., a wholly-owned subsidiary of Union Oil Company of California. Small scale underground mining first began in 1920 and large scale open-pit mining was initiated in 1965, but because of the high stripping ratio, surface mining is being phased out. Construction of the underground mine began in 1979, and first production is scheduled for 1983. Full production of 9 Gg of molybdenum per year will be reached in 1984. Access to the mine will be through a service shaft, development shaft, and a conveyor decline. &apos;The service shaft, 7.3 m in diameter and 396 m deep, will be used for personnel, supplies, development muck, and fresh air intake. It will be equipped with a large, double-deck man and material cage. A double-drum hoist will be used for cage hoisting and will handle up to a 27-t load in balance with a counterweight. Development muck will be loaded into skips in balance and a single-drum auxiliary hoist will provide secondary access. The 396-m-deep, 4.3-m-diameter development shaft will be used for initial mine development, mine exhaust air, and a slick line pass for concrete entering the mine. The development shaft is located 400 m north of the service shaft. Run-of-mine ore will be transported from one of two 1.4-kt storage pockets by a 1.2-m-wide conveyor belt located in the 10°, 1980-m-long, 5.0-m-wide, and 4.6-m-high decline. The decline conveyor belt has a centerline length of 2229 m and a vertical rise of 363 m. It will be powered by four 448 kw motors driving a single drive pulley. There will be two 1.2 m transfer belts on the surface with centerline lengths of 322 m and 542 m. All three conveyor belts will be reinforced with steel cables. On the surface, in addition to the headframes and hoist houses for the two vertical shafts, a large maintenance shop, warehouse, compressor house, change house, main office building and timber framing building have been constructed.

Jan 1, 1981

By H. H. Moss

OXYACETYLENE .cutting has experienced rapid development in the last few years and greater advances and expansion and broader application may be expected in the immediate future. Marked changes in cutting practice may be expected as a result of the advances in metallurgical science that have produced special alloy steels with properties requiring different cutting techniques than for ordinary low-carbon steels. Much current attention is being directed to those problems of modified cutting practices. Cutting has thus ceased to be merely a shaping problem. It is rapidly coming to have a predominantly metallurgical aspect. The problem is no longer how to cut a given shape, but how to cut a given steel. The answer

Jan 1, 1936

By A. A. Gustafson

The Freeport Sulphur Co. built a portable crusher in 1944 that solved a problem that none of the crusher manufacturing companies contacted were able to do. Two of these crushers have now been in service for 6 years, giving satisfactory service in handling a million tons per year, and having a capacity of over 2000 long tons per 8-hr shift. These machines reduce ton size blocks of sulphur to 12 in. and also feed the conveyor system at the rate of 350 long tons per hr per machine.

Jan 10, 1950

By H. C. Meister

THE zinc-smelting industry of the United States has grown very rapidly in recent years and bids fair to outrival that of all other countries in the future. On account of the geographical situation of the chief ore-supplies, the industry may be divided into two main groups, the eastern and the western. The eastern and older group includes the works of Pennsylvania, New Jersey and Virginia, which practically control their own ore-supply. The western group includes the works of Illinois, Missouri, Kansas, Wisconsin and Indiana, which do not control their own ore-supply, but are compelled to buy the ore in the open market. It is for this reason that the western zinc-smelting industry has been undergoing continual changes since its inception 40 years ago. During the early discoveries of lead- and zinc-ore in the lead- and zinc-districts of Wisconsin and northern Illinois, the zinc-ore (sinithsonite or " bone " as the miner calls it) had no value and was considered waste; but later, when this ore was found in large quantities, it attracted the attention of metallurgists, which finally resulted, during the early sixties, in the building of the Matthiessen & Hegeler Zinc Works at LaSalle, Ill., the pioneer zinc-works of the West. A few years later, zinc-ore was discovered associated with the lead-ore in Jefferson county, southeastern Missouri, which started the industry at Carondelet or South St. Louis. Still later, the discoveries of zinc-ore in the lead-ore region, first&apos; at Granby and then around Joplin, lead to the introduction of the industry in the southwest Missouri and Kansas districts. At first the development of the industry was slow, but it progressed steadily and it is now one of the leading metallurgical industries of the middle West.

Jan 1, 1905

By T. S. Taylor

THE development of the linear hot-wire anemometer has been chiefly clue to the efforts of L. V. Kings1 and A. E. Kennelly and H. S. Sanborn.2 The anemometers used by these investigators consisted essentially of a fine heating wire having attached leads for resistance measurements at distances of 10 or more centimeters from each other. In using such an anemometer, the current is measured that is necessary to maintain the resistance of the wire, between the two leads, constant for different air velocities. Thus resistance is always so chosen that the temperature of the heating wire will be sufficiently high to make the observations practically independent of small variations in the temperature of the gas in which the anemometer is placed. The measurement of the resistance of the anemometer wire requires a Kelvin bridge set up, which for commercial work is not altogether desirable. The cooling effect clue to different air velocities depends on the temperature difference between the wire and the gas and the total quantity or mass of gas passing the wire per unit tine. Since the temperature of the wire is maintained constant, the effect observed in the change in the heating currentfor a given change in air velocity is a measure of the difference between the gas flow in the two cases. Such an anemometer, therefore, measures the average gas flow for a length depending on the distance between the two resistance leads. Therefore the instrument in the form used thus far is not very satisfactory for measuring gas velocities in small tubes or in places where the velocity varies rapidly across the line of flow. For the purpose of measuring gas flow through relatively small tubes, where the velocity changes rather rapidly across the tube, the hot-wire anemometer has been modified in the manner shown diagrammatically in Fig. 1. It consists essentially of a platinum heating wire H about 0.007 in. (0.178 mm.) in diameter and ½ to 1 in. (12.7 to 25.4 mm.) long stretched across a suitable framework, say of glass. This wire has

Jan 9, 1919

By Grafton M. Thrasher

For the proper discussion of this subject it is necessary to incorporate in this paper the substance of part of an article published by me in Metallurgical and Chemical Engineering for January, 1915. Several years ago, while I was chief chemist for a malleable-iron company, it occurred to me that the mere regulation of silicon was not sufficient to produce the desired degree of chill. But only after repeatedly encountering this fact in both open-hearth and air-furnace, as well as cupola metal, was I led to regard variation in the total carbon as the principal cause of disturbance; and I resolved to determine the total carbon as well as the silicon in every heat, and by a carefully prepared test piece to gauge the approximate degree of chill. The test piece selected was a round plug, 2 in. in diameter and 12 in. long, cast on end

Jan 1, 1916

By John C. McDonald

This report is a continuation of an earlier one with a similar titlell to which the reader is referred for such details of procedure as do not appear here. A brief summary will be given of the objects and methods of the research at this time. It was pointed out in the previous paper that properties of magnesium alloys, as reported by different workers, often differ. This was thought to be due to differences in the preparation of the specimens tested. The ideal way to compare alloys would be to prepare each of them in so many ways that reasonable assurance would exist that the best properties of each alloy had been developed. The alloys would then be compared on the basis of these properties. The treatment actually given the alloys of these studies is necessarily a compromise with completeness. Nevertheless, enough treatments have been given so that a considerable degree of assurance exists that the properties obtained are representative of the alloy as such, and not merely of the manner in which it was worked and heat-treated. Procedure The alloys were cast in the form of cylindrical ingots, which were extruded into bar before being rolled. Some of the ingots were made in 2-in. diameter molds whereas all of the alloys of the previous paper were cast in 4.in. molds. For the larger ingots, the extruded bar was 2 by 0.140 in.; lor the 2-in. ingots, it was I by 0.100 in. Preliminary work indicated that strip rolled from both sizes of the bars would have substantially the same properties. The bar was rolled at some temperature in the range 400&apos; to 800°F., for a total reduction of 50 per cent. The temperature and reduction per pass were so selected that no cold cracking occurred, but were such as to leave the metal in a semihard condition. The hot-rolled metal was annealed at several temperatures in the range 400&apos; to 800°F.; it was also cold-rolled, and given solution and aging treatments. Two standard test bars corresponding to each treatment were pulled in tension, and the average was obtained. Twelve treatments, in all, were given each alloy. The properties of the annealed specimen having the highest percentage of elongation were selected to represent the alloy in the annealed condition. The properties of the cold-rolled specimen having the highest yield strength were selected to represent the alloy in the cold-rolled condition. These properties were then plotted against percentage composition by analysis. Smooth curves were drawn among the points, thus graphically averaging the results over the entire range studied. The percentage of elongation in the cold-rolled state was not plotted, since it was always low, and varied in an irregular manner with composition. Alloys Studied Magnesium-calcium Alloys.—Fig. I presents the results on the alloys of the rnag-

Jan 1, 1941

By John C. McDonald

This report is a continuation of an earlier one with a similar titlell to which the reader is referred for such details of procedure as do not appear here. A brief summary will be given of the objects and methods of the research at this time. It was pointed out in the previous paper that properties of magnesium alloys, as reported by different workers, often differ. This was thought to be due to differences in the preparation of the specimens tested. The ideal way to compare alloys would be to prepare each of them in so many ways that reasonable assurance would exist that the best properties of each alloy had been developed. The alloys would then be compared on the basis of these properties. The treatment actually given the alloys of these studies is necessarily a compromise with completeness. Nevertheless, enough treatments have been given so that a considerable degree of assurance exists that the properties obtained are representative of the alloy as such, and not merely of the manner in which it was worked and heat-treated. Procedure The alloys were cast in the form of cylindrical ingots, which were extruded into bar before being rolled. Some of the ingots were made in 2-in. diameter molds whereas all of the alloys of the previous paper were cast in 4.in. molds. For the larger ingots, the extruded bar was 2 by 0.140 in.; lor the 2-in. ingots, it was I by 0.100 in. Preliminary work indicated that strip rolled from both sizes of the bars would have substantially the same properties. The bar was rolled at some temperature in the range 400&apos; to 800°F., for a total reduction of 50 per cent. The temperature and reduction per pass were so selected that no cold cracking occurred, but were such as to leave the metal in a semihard condition. The hot-rolled metal was annealed at several temperatures in the range 400&apos; to 800°F.; it was also cold-rolled, and given solution and aging treatments. Two standard test bars corresponding to each treatment were pulled in tension, and the average was obtained. Twelve treatments, in all, were given each alloy. The properties of the annealed specimen having the highest percentage of elongation were selected to represent the alloy in the annealed condition. The properties of the cold-rolled specimen having the highest yield strength were selected to represent the alloy in the cold-rolled condition. These properties were then plotted against percentage composition by analysis. Smooth curves were drawn among the points, thus graphically averaging the results over the entire range studied. The percentage of elongation in the cold-rolled state was not plotted, since it was always low, and varied in an irregular manner with composition. Alloys Studied Magnesium-calcium Alloys.—Fig. I presents the results on the alloys of the rnag-

Jan 1, 1941