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By James L. Poole

INTRODUCTION Companies in the mining industry are subjected at times to currency exchange rate risks and interest rate risks. The former occurs any time a firm deals in more than one currency. The latter occurs anytime a firm is financing itself with borrowings that have a floating interest rate (although a fixed rate can be a problem at times). FOREIGN EXCHANGE RISKS There are two general reasons why exchange rate risks (FX risk) can occur: First, a firm may be building a mine or plant in another country where financing for the plan is denominated in funds other than the country hosting the plant. For example, if a Canadian firm is building a smelter in another country which is being financed with Canadian dollars, the C$100 million for labor, which must be paid in the local currency, has subjected the firm to a C$100 million FX exposure risk. If after the cost estimates were made the local currency appreciates 30% against the Canadian dollar, the cost of labor has increased to C$130 million. (Of course, the local currency could also devalue during this period making the labor cost component less than estimated.) The other reason exchange rate risk arises is due to the fact that the costs of production in one country producing minerals may be incurred in one currency while the production is sold in another currency, such as coal being exported from Canada to Japan under contracts denominated in Yen. In this example, the Canadian firm is exporting coal to Japan and being paid in yen in 30 days at a contract price of 100 Y per ton. Furthermore, assume that the mining firm anticipates a depreciation of the yen against the Canadian dollar of 5% in the next 30 days. Therefore, the mining firm stands to lose 5% of the value of its shipment in a month's time not including carrying costs. In both cases, the mining company risks an adverse change in the exchange rate, that is, the currency the company holds may decrease in value relative to the other currency. There are a number of ways to mitigate the exposure to foreign exchange which can be utilized by the mining company. Optimally, a company should try to match all their inflows and all their outflows in the same currency. For example, if a Canadian firm has mining costs and debt service costs both denominated in Canadian dollars then 100% of their receipts would need to be in Canadian dollars. If a Canadian firm has 70% of its cash outflows in Canadian dollars and 30% of its cash outflow in American dollars in the form of a loan amortization then their receipts would need to be 70% in Canadian dollars and 30% in U.S. dollars. This general principle of matching cash receipts and cash costs in the same currency works fine in a perfect world but what of the real world when these cannot be matched? In actual practice, however, the financial markets have developed a number of ways in which foreign exchange risks can he managed or reduced. FORWARD MARKETS One method is to use the forward markets which are made on a world-wide basis by the commerical banks for foreign exchange. If a firm wanted to lock in the current exchange rate for the receipt of a foreign currency at some point in the future, they could do so by contracting on the forward market with a bank selling that foreign currency at the current rate at the future date. For example, assume that a Canadian firm is exporting coal to Japan which would pay yen in 90 days. If a firm wanted to lock in today's exchange rate they could contract with a bank to sell on the forward market yen in 90 days. When the yen was received the contract would be executed by selling the yen to the bank and receiving the previously agreed upon number of Canadian dollars. Generally speaking, the forward market can be used to sell forward about 12 months. The costs of selling (or buying) forward

Jan 1, 1985

By Robert Williams

A description is given of a double heat treatment of 60-40 brass. Photomicrographs are included to show the changes that take place in the microstructure on reheating the water-quenched specimens. A wide variation in the mechanical properties may be obtained by the described heat-treatment method. ON SEVERAL occasions, when 60-40 brass is first obtained in the beta condition by quenching at about 825° C. and is then reheated, the writers have noticed that recrystallization will take place in the form of twin crystals at the grain boundaries of the original beta crystals. It was, therefore, thought that interesting results would be obtained if this twinning process was studied further, especially as to the temperature at which it first occurs. The material used in the investigation was a bar of extruded muntz metal, 1/2 in. in diameter, and of the following composition: Copper, 60.21 per cent.; zinc, 39.73 per cent.; lead, trace; iron, trace; tin, nil. Specimens I in. in length were sawed from this bar and a hole was drilled into the side of each specimen until it reached the center, and the tip of a platinum-platinum rhodium thermocouple was inserted. A small Hoskins' resistance-type electric furnace was used for heating the specimens. The specimens were first heated at 840°C. for 2 hr. and then quenched in cold water, in order to get the material entirely in the beta state. In every case, in the reheating process, the furnace was brought to the desired temperature before the specimen was introduced. The time occupied in preheating was from 20 to 30 min. By sawing back 1/4 in. on each side of the hole, two parts were obtained from each specimen

Jan 3, 1924

(The following list contains the names of those members of the Institute of whose connection with military service we have only recently become acquainted; it also includes the names of a few who have recently been promoted or transferred, indicated by a *. A complete list was published in the Year Book, issued as a supplement of the Bulletin for March, 1915.) ADAMS, ARTHUR K., Lieut., Co. 2, E. 0. T. S., Camp Humphreys, Va. AMIDON, CLAUDE E., 11th Training Co., Casual Camp, Camp Cody, N. M *BARBOUR, P. E., Capt., Engineer Corps, U. S. A. *BENHAM, W. M., Master Engr., Headquarters Dept., 115th Engrs. A. E. F., Care Postmaster, N. Y. BICKNELL, H. L., Pvt., Co. D, 27th Engrs., Camp Leach, D. C. *BLANDY, S. H. B., Lieut., Royal Horse Artillery, B. E. F.

Jan 10, 1918

By Benjamin F. Tillson

THE past seven years, and 1937 in particular, have witnessed the return of New Jersey iron mining to a place of importance. Following the World War period, little mining was done for several reasons. The properties were operated by small merchant furnaces whose demand was not sufficiently large to warrant large-scale methods with their attendant low operating costs. Under those conditions, the ore produced could not compete with that from the Great Lakes. However, when on Nov. 1, 1.929, the Alan Wood Steel Co., of Conshohocken, Pa., leased the Scrub Oak and Washington properties at Mine Hill and Oxford respectively, they made radical changes in the mining and milling processes which greatly reduced the cost of production and pointed the way to a renewed interest in mining in New Jersey. The following factors have been advanced as influencing the reopening of New Jersey mines:

Jan 1, 1938

By P. Somasundaran, I. J. Lin

There are several methods described in the literature for the determination of the order and the rate constant for the flotation of minerals. These often involve some type of computational or graphical procedures. A relatively simpler method developed on the basis of a technique used by Chang" in chemical kinetics is described in this paper. Chang's technique, like most others in chemical kinetics," is based on the law of mass action which states that the rate of a chemical reaction is proportional to the active masses of the reacting substances at any given time. Application of this law to the flotation of minerals would involve the adaptation of the terms "flotation" for "chemical reaction" and "floatable amount of the minerals" for "the active masses of the chemical reactants." There are indeed a large number of physical processes controlling the actual flotation rate. The most inefficient one of these could be considered equivalent to the slowest reaction in a series of chemical reactions. If variables such as aeration rate and agitation are kept at constant levels, the rate of flotation under free flotation conditions (i.e., conditions under which the suspension is dilute) will be given by the following equation: where C is the concentration of the floatable minerals in the pulp at any given time t, K is the rate constant, and n is the order of the reaction. If we represent the floated fraction of the floatable mineral at time t as m (0 m - 1) and C at t = 0 as Ca, integration of Eq. 1 yields

Jan 1, 1974

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

YOUNG'S modulus of elasticity in metals ordinarily decreases with rising temperature. The range of the thermoelastic coefficient at room tem- 1/E dE/dT perature (temperature coefficient of modulus, is illustrated1 by the values for tungsten and hard-drawn aluminum, —95 and —1220x10-6 eg C-1. However, a nearly zero thermoelastic coefficient is observed in some ferromagnetic alloys over a limited temperature range. In these alloys, as the temperature is raised, an increase in modulus arising from loss of ferromagnetism compensates the normal decrease. Iron-nickel alloys containing 36 to 52 pct Ni behave in this manner. In the present investiga- tion, the dependence of the modulus and thermo-elastic coefficient on cold working and annealing were studied in iron-nickel alloys in this range of nickel. Although these alloys, when fully annealed, have a coefficient near zero over a very small temperature interval, the temperature range of low coefficient can be extended greatly by straining and then stress-relief annealing. The modulus of face-centered cubic iron-nickel alloys annealed at 1000°C does not vary linearly with composition at room temperature but, as shown by the solid line of fig. 1, from Koster,' is minimum at 40 pct Ni and maximum at 87 pct Ni. The modulus increases with applied magnetic field but retains the minimum near 40 pct Ni. Schematic change of modulus with temperature for three compositions is shown in fig. 2.'. ' With increasing temperature, the slope becomes less negative, goes through zero (temperature of minimum modulus for the alloy), becomes positive, and resumes its normal negative value above the Curie

Jan 1, 1951

[Arizona BRENT N. RICHARD, Chairman ROBERT W. HUGHES, Vice-Chairman ROBERT E. S. HEINEMAN, Secretary-Treasurer Arizona Bureau of Mines, University of Arizona, Tucson, Ariz. THOMAS G. CHAPMAN A. MENDELSOHN WILLIAM KOERNER A. C. REIFSNIDRR C. R_ . KUZELL WILLIAM M. SABEN ERNEST WITTENAU Black Hills NATHANIEL HERZ, Chairman C. E. DAWSON, First Vice-Chairman BANCROFT GORE, Second Vice-Chairman C. E. SCHMIDT, Secretary-Treasurer Homestake Mining Co., Lead, S. D. R. S. BLITZ NATHANIEL HERZ C. E. DAWSON A. I. JOHNSON L. B. EAMES F. C. LINCOLN BANCROFT GORE C. E. SCHMIDT]

Jan 1, 1940

By Robert C. Stephenson

Large deposits of titaniferous magnetite occur associated with anorthosite and gabbro in the Lake Sanford area, Essex County, New York. The ore, gabbro, and anortbosite show consanguineous relations. Anorthosite grades through gabbroic anortbosite into gabbro but is locally intruded by gabbro. Structural and petrographic evidence suggests that gabbro was differentiated from anorthosite by filter pressing. Field and laboratory study of ore occurrences indicates that an ore residuum concentrated in the gabbro through magmatic segregation' by normal crystallization and flowage. Ore occurs as bands which grade into gabbro, and also as large masses which replace anorthosite. Ore in the gabbro is finer grained and richer in ilmenite than ore in anorthosite. Introduction The Lake Sanford area lies in the heart of the Adirondack Mountains, in the southeast portion of the Santanoni Quadrangle (Fig. I), and there are four ore bodies in the area immediately surrounding Lake Sanford. The area lies 30 miles by highway from the railroad at North Creek. The deposits were discovered in 1826, and from then until 1858 iron was produced in small quantities. They were idle from 1858 to 1906. Between 1906 and 1913, extensive development work was carried on. These early attempts to produce iron ore were thwarted by transportation difficulties and by the titanium content of the ore, which, ironically, has been the element responsible for the present development. In 1941 the National Lead Co., Titanium Division, developed the Sanford Hill ore body and is now producing ilmenite and magnetite concentrates.18 The Defense Plant Corporation is extending the railroad to the property from North Creek and is building a sintering plant to treat the magnetite concentrate. Investigations OF the Deposits* Early studies of the Lake Sanford titaniferous magnetite deposits were made by Emmons in 1842,l2 by Kemp, in 1897-98'0 and by Newland in 1908.22 Singewald26 and Osborne23 considered these deposits along with numerous other titaniferous magnetite deposits in 1913 and 1928, respectively. None of these reports entailed a detailed study of the ore occurrence in this area. The present study was begun in 1940. Seven months were spent in the field, and considerable time was spent in labora-

Jan 1, 1949

By Robert C. Stephenson

Large deposits of titaniferous magnetite occur associated with anorthosite and gabbro in the Lake Sanford area, Essex County, New York. The ore, gabbro, and anortbosite show consanguineous relations. Anorthosite grades through gabbroic anortbosite into gabbro but is locally intruded by gabbro. Structural and petrographic evidence suggests that gabbro was differentiated from anorthosite by filter pressing. Field and laboratory study of ore occurrences indicates that an ore residuum concentrated in the gabbro through magmatic segregation' by normal crystallization and flowage. Ore occurs as bands which grade into gabbro, and also as large masses which replace anorthosite. Ore in the gabbro is finer grained and richer in ilmenite than ore in anorthosite. Introduction The Lake Sanford area lies in the heart of the Adirondack Mountains, in the southeast portion of the Santanoni Quadrangle (Fig. I), and there are four ore bodies in the area immediately surrounding Lake Sanford. The area lies 30 miles by highway from the railroad at North Creek. The deposits were discovered in 1826, and from then until 1858 iron was produced in small quantities. They were idle from 1858 to 1906. Between 1906 and 1913, extensive development work was carried on. These early attempts to produce iron ore were thwarted by transportation difficulties and by the titanium content of the ore, which, ironically, has been the element responsible for the present development. In 1941 the National Lead Co., Titanium Division, developed the Sanford Hill ore body and is now producing ilmenite and magnetite concentrates.18 The Defense Plant Corporation is extending the railroad to the property from North Creek and is building a sintering plant to treat the magnetite concentrate. Investigations OF the Deposits* Early studies of the Lake Sanford titaniferous magnetite deposits were made by Emmons in 1842,l2 by Kemp, in 1897-98'0 and by Newland in 1908.22 Singewald26 and Osborne23 considered these deposits along with numerous other titaniferous magnetite deposits in 1913 and 1928, respectively. None of these reports entailed a detailed study of the ore occurrence in this area. The present study was begun in 1940. Seven months were spent in the field, and considerable time was spent in labora-

Jan 1, 1949

By R. M. Lumley, J. D. Wood

In this investigation, polyethylene terephthalate (Mylar) was coated with various thicknesses of a vapor-deposited silvel- precoat followed by a uapor-deposited zinc using several orders of- magnitude variation in deposition rate of both the silver and zinc. The grain size and resistivity of the zinc film was primarily a function of the silver precoat thickness (in the range of 0.001 to 10 average Agprecoat thickness) and a secondary function of the zinc film thick-ness. The silver and zinc deposition rates were found to have a minor effect on the characteristics of- the zinc film. Electron and X-ray dilffraction results slwwed that all zinc films exhibited a fiber texture mith the (0001) axes essentially perpendiculav to the platze of the film When coating nonmetallic substrates such as paper or plastics with zinc metal vapors, it is found that the zinc will not condense unless extremely high evaporation rates or very low substrate temperatures are used. In 1931, Fraser' reported a procedure whereby zinc could be made to condense on nonmetallic substrates by depositing a very thin precoat of silver. Utilizing Fraser's method, the ermans' developed a commercial process of making "metallized film" capacitors during World War 11. The metallizing process, which was introduced in the United States shortly after the conclusion of World War 11, became one of the first large-scale commercial applications of vapor-deposited metal films and remains one of the larger applications in the electronics industry. The process, as developed by the Germans, consisted of depositing a thin precoat of silver on lacquered, kraft paper followed by a conducting layer of zinc. The resulting silver-zinc film was used as the capacitor The term silver-zinc will be used throughout thiz paper to mean a metal film formed by an extremely thi layer of silver, 10A or less in average thickness, and a zinc film IOOA or more in thickness deposited on the silver film. ---. electrode with the lacquered paper being the dielectric. When polyethylene terephthalate was marketed by E. I. Du Pont Co. in the early 1950's under the trade name Mylar it was substituted for the kraft paper in many capacitor applications. Even though manufacture of metallized paper and metallized Mylar capacitors has been a large commercial process for many years, little information has been published about the deposition parameters. Correlating the work of others to the various effects to be expected of deposition parameters on silver-zinc films deposited on Mylar, it is expected that the

Jan 1, 1967

By Edward H. Robie

IN many ways, the Annual Meeting of the Institute in February was the most successful yet held. Attendance surpassed even that of the 75th Anniversary Meeting in 1947 with its international flavor. The number of technical sessions and the number and quality of papers presented almost certainly set a new record. The banquet was the best attended of any yet held even though the price reached a new high. Despite the demand for almost 1700 tickets to the banquet, some felt that the price-$13.50-was higher than they could afford. There are two points of view: One is that this is the big social event of the year for members and that it is worth from three to five dollars more to have it at the Waldorf and to have a luxury menu, as we did this year, with turtle soup, a fish course, guinea hen for the entree, and wild rice (which we understand is $3.50 per lb) for a vegetable. The other point of view is that a seven or eight dollar banquet of more simple food at the Commodore Hotel would more than make up in appeal, especially to the younger and less well-heeled members, what it would lack in glamour.

Jan 1, 1952

By AIME AIME

INDUSTRIAL Pittsburgh, the center of the coal and iron and steel industry of the world, was host to the Coal Division at its Fall Meeting held there on Oct. 21 and 22 at the William Penn Hotel. The past year has witnessed remarkable progress in new methods of mining and burning of coal and many of the papers on the program discussed these developments. Nearly 100 had registered at the time of the first session on Wednesday morning and this number increased to 168 on Thursday. Owing to the large number of papers on the program, eighteen in all, it was necessary to hold group meetings on Thursday morning, with a capacity attendance at each meeting. . Chairman McAuliffe was on hand early Wednesday morning to open the meeting. George Otis Smith, representing President Lovejoy, came over from Cleveland where he attended the meetings of the Institute of Metals Division. Secretary Parsons arrived later from Cleveland and spent the whole morning with the executive committee of the Division. Doctor Smith, Secretary Parsons, and J. T. Breunich were the official representatives from Institute headquarters.

Jan 1, 1936

By C. H. Wilson, J. J. Jakosky

In recent years geophysical prospecting methods have become well established as important steps in economically initiating new mining ventures in the field of base-metal exploration, placer mining and water-supply problems. The value of these methods as applied to these problems has been greatly increased by new and improved instruments, and a greater degree of skill in laboratory and interpretative technique. Considerable success has been attained, with subsequent drilling checking the predicted depths within 1 to 5 per cent, on problems requiring information on thickness of fill material, depth to and contour of bedrock, extent and yardage of gravels to be mined, and location of hydrostatic water table. In these applications the geophysical work gives data comparable to drilling, and at a cost of about one-fiftieth that of drilling. Discussion of several surveys with accompanying illustrations showing geophysical methods applicable to both simple and complicated placer problems are contained in the paper. The electrical instruments for measuring electromagnetic field strength and surface potentials are described, and their use, coupled with that of other geophysical instruments in modern field procedure, is noted.:

Jan 1, 1934

United Engineering Trustees, Inc James L Head, '55, E C Meagher, '56, Gail F Moul¬ton, '57 Advisory Committee to Real Estate Committee Cleveland Dodge, Herbert Hoover, Paul D Merica The Engineering Foundation W M Peirce, '56, A M Gaudin, '58 Research Procedure Committee A M Gaudin, March '54 Engineering Societies Library Board Frank T Sisco, '54, E M Wise, '56, James L Head, ex officio Engineering Societies Personnel Service Board of Directors E J Kennedy, Jr, E H Robie

Jan 1, 1954

By P. E. Busby, C. Wells

FURTHER study of data used in determinations of 1—rates of diffusion of boron in austenite and 2—solubilities of boron in the a and phases of iron and steel' has provided an equation for the diffusion of boron in a iron. In brief, the previously published data were accumulated from the results of deboron-izing (and decarburizing) experiments carried out in the range of 700° to 1300°C. Diffusion coefficients (D?) for boron in austenite were calculated using the Grube solution of Fick's law. However, only solubility values were estimated from the discontinuous concentration-penetration curves, Fig. 1, which are characteristic of diffusion in two phases. Dr. Carl Wagner' has recently provided a solution for calculating D values from penetration curves of this type as follows: Cl1.1 —C8/C8- C11.1 = vp ? e ? erf (?) and D = ?/4?2 t where C11,12, C8, C11, and E have the values shown in Fig. 1. D is the diffusion coefficient, sq cm per sec; t is time of deboronizing anneal, see; and ? is a di-mensionless parameter. For a given diffusion experiment, the value of ? can be readily obtained by graphical solution from a plot of y vs the right-hand part of Eq. 1. D may then be evaluated from Eq. 2. The avplication of this solution to previously re-ported results, of which excerpts are given in Table I, permits the calculation of diffusion coefficients for boron in a iron. On the basis of these meager data, it is tentatively concluded that the diffusion of boron in a may be represented by the equation Da = 1011 e-92,900/RT where R is the gas constant, cal X 0C-1 X mol-1; and T is the absolute temperature, OK. Although the frequency factor, 106 sq cm per sec, is admittedly several orders of magnitude higher than expected, the value of Q, 62,000 cal per mol, appears reasonable and is, in fact, very similar to that for the self-diffusion of iron. It is pertinent to mention at this point that the value of Q obtained for the diffusion of boron in austenite by means of the Wagner solution is 19,000 cal per mol and is in excellent agreement with the value previously reported1 in the equation DT = 2xl0-3 e-21,000/RT [4] which was determined by the application of the Grube solution to other data. The fact that determined constants A and Q in the equation D? = A e-Q/RT were practically the same, independent of whether the Wagner or Grube solutions were used in determinations of D values, strengthened the authors' belief that the computed values of D (Table I) using the Wagner solution are reliable. The relative values in Eqs. 3 and 4 for the diffusion of boron in the a and ? phases, respectively, suggest that boron forms a substitutional solid solution in a iron and an interstitial solid solution in ? iron. The same tentative conclusion has been reached by McBride et al.3 on the basis of relative solubilities, atom diameter, and the size of the interstitial hole in a and ? iron. In connection with the data for test 14 listed in Table I, it is of interest to calculate the solubility of boron in a iron using the D value given by Eq. 3. As might be anticipated from the small movement of the interface in test 14, proper substitutions in Eqs. 2 and 1 give a low value, approximately 0.0004 pct B, at 850°C. Apparently at 835°C (test 13) it is possible to obtain 0.0018 pct B, and at 850°C only 0.0004 pct B into solution in the a phase before a second phase appears. These observations are consistent with the Fe-B constitution diagram proposed by McBride, Spretnak, and Speiser.3 References P. E. Busby, M. E. Warga, and C. Wells: Diffusion and Solubility of Boron in Iron and Steel. Trans. AIME (1953) 197, p. 1463; Journal of Metals (November 1953). 'W. Jost: Diffusion in Solids, Liquids, Gases. (1952) pp. 69-71. New York. Academic Press Inc. "C. C. McBride, J. W. Spretnak, and R. Speiser: A Study of the Fe-Fe2B System. Trans. ASM (1954) 46, p. 499.

Jan 1, 1955

ROBERT H. LEACH, Chairman ROBERT F. MEHL, Past-Chairman D. K. CRAMPTON, Vice-Chairman E. M. WISE. Vice-Chairman W. M. CORSE, Treasurer LOUIS JORDAN, Secretary 29 West' 39th St., New York, N. Y. Executive Committee 1940 1941 ARTHUR PHILLIPS W. P. SYKES L. E. SCHUMACHER E. A. ANDERSON CARL E. SWARTZ D. L. COLWELL 1942

Jan 1, 1939

By R. W. Raymond

THE death of Privy Councilor Bruno Kerl, on March 25, 1905, terminated a distinguished and useful career. Bruno Kerl was born March 24, 1824, at St. Andreasberg in the Harz, and entered in 1840 the mining school of Clausthal, where he studied for three years, working at the same time in the mines and stamp-mills. After an additional year of practice in the silver-reduction and iron-works of the Upper Harz, he passed his examination as a metallurgical graduate, and studied for a year at Göttingen under WV filer in chemistry, Hausmann in mineralogy, and eminent teachers in other branches. In 1846 he was employed at the Oker works, but very soon left this post, to become a tutor in chemistry, metallurgy, etc., at Clausthal, where he continued to labor until 1867. During this period the Clausthal school became a completely equipped Mining Academy, and Kerl was made in 1862 a full professor. Meanwhile, in addition to the work of instruction, he discharged, simultaneously or successively, the duties of many responsible official positions in the Royal Hanoverian administration of mines and smelting-works, and also published many valuable books and treatises. Among these were professional guide-books to the Upper and Lower Harz, and their characteristic metallurgical processes; a manual of qualitative and quantitative blow-pipe analysis; another of metallurgical assaying ; and, most important of all, a Handbook of Metallurgy in four volumes, the second edition of which was issued in Leipzig, 1861-1865. This book was speedily recognized as the leading authority in its field. In this country, the English translation and revision, by Crookes and Röhrig, published about 1869, had a large circulation. In addition to these literary labors, Prof. Kerl became in

Jul 1, 1905

Trans No. Place Date Vol. 1. Wilkes-Barre, Pa' May, '71 1 2. Bethlehem, Pa Aug., '71 1 3. Troy, N. Y Nov., '71.. 1 4. Philadelphia, Pa Feb., '72.. 1 5. New York, N. Y.' May, '72.. 1 8. Pittsburgh, Pa Oct., '72.. 1 7. Boston, Maas Feb., '73.. 1 8. Philadelphia, Pa." May, '73.. 2 9. Easton, Pa Oct., '73.. 2 10. New York, N. Y Feb.. '74.. 2 11. St. Louis, Mo.* May, '74.. 3 12. Hazleton, Pa Oct., '74.. 3 13. New Haven, Conn Feb., '75.. 3 14. Dover, N. J. ? May, '75.. 4 15. Cleveland, 0 Oct., '75.. 4 16. Washington, D. C Feb., '76.. 4 17. Philadelphia, Pa.} June, '76.. 5 18. Philadelphia, Pa Oct., '76.. 5 19. New York, N. Y Feb., '77.. 5 20. Wilkes-Barre, Pa. ? May, '77.. 8 21. Amenia, N. Y Oct., '77, . 6 22. Philadelphia, Pa Feb., '78:. 6 23. Chattanooga, Tenn.*...May, '78'.. 7 24. Lake George, N. Y Oct., '78.. 7 25. Baltimore, Md.* Feb., '79.. 7 26. Pittsburgh, Pa Mayo '79.. 8 27. Montreal, Canada Sept., '79., 8 28, New York, N. Y.? Feb., 'SO.. 8 29. Lake Superior, Mich. ..Aug., '80.. 9 30. Philadelphia, Pa.' Feb., '81.. 9 31. Staunton, Va May, '81..10 32. Harrisburg, Pa Oct., '81..10 33. Washington, D. C.? ....Feb., '82..10 34. Denver, Colo Aug., '82..11

Jan 1, 1928

By Werner Schwartz, Wolfgang Haase

Lead sulfide concentrates, which may include other lead concentrates, are sintered on an up-draught sintering machine without the addition of any diluting agents or fluxes. Subsequently they are melted in an oil- or gas-fired rotary furnace. The sintering and melting processes are based upon the following roast-reaction: PbS + 2 PbO = 3 Pb + SO, PbS + PbSO, =2 Pb + 2 SO, For obtaining a lead bullion free from sulfur, the sintering process is carried out in such a way that the sinter product contains a small amount of excess oxygen above that to react with the sulfides. At the end of the melting process, when the reactions are finished, the remaining small amount of oxide residues is reduced with coal to which a certain percentage of soda ash (about 1 pct of the lead bullion) is added. For the lead smelting process described neither coke nor fluxes—except soda ash—are required. This process is being utilized by a European smelter successfully and with a high lead recovery. The consumption figures for the smelting of 100 tons per day of lead concentrates are indicated. The lead content of the lead concentrates from modern ore dressing plants ranges from 65 pct to above 80 pct. In most lead smelters of the world these concentrates are smelted in a blast furnace. For blast-furnace smelting the concentrates have to be desulfurized and agglomerated by sintering. A requirement for the perfect operation of a down-draught sintering machine and of a blast furnace is a maximum lead content in the feed of 40 to 45 pct. For this reason, some lead concentrates have to be diluted by adding return slags, limestone, and possibly iron oxide and sand. As an example, 100 tons of lead concentrate with 72 pct Pb would contain 13.5 tons of gangue (including the zinc). To produce a perfect sinter with 42 pct Pb it would be necessary to add 70 tons of flux and return slag, more than five times the original weight of the gangue, to the sinter mix and blast-furnace charge. A correspondingly large amount of coke would be required in order that all of these materials reach the heat of formation and the melting temperatures of the slag (1200" to 1400°C) inside the blast furnace. The roast-reaction process presents a possibility for lead recovery without dilution of the concentrates. In this process the concentrate mixed with coal is placed upon a Newnam-hearth and air is blown through nozzles into the heated mix. AS a result metalllic lead and a relatively great amount of so-called .'Grey Slag" with a lead content of 25 to 35 pct are formed. The slag is sintered to eliminate sulfur and, after addition of the requisite fluxes, treatt:d in a blast furnace. Owing to the poor recovery of lead from the hearths and to the unavoidable heavy hand-work plus the risk of poisoning this process is utilized in very few 112ad smelters today. Since in mxny countries of the world coke is expensive and difficult to obtain, it appeared feasible to use the principle of the roast-reaction by modern sintering and melting methods with recovery of the lead in electric, or oil, gas, or coal-fired furnaces. Two processes are utilized on an industrial scale: A) Lead smelting in the electric furnace of the Bolidens Gruv A/B in Sweden, as described by S. J. Walldcn, N. E. Lindvall, K.G. Gorling, and S. Lundquist. B) The self-fluxing lead smelting of Lurgi Gesell-schaft fiir Chemie und Huttenwesen m.b. H., Frankfurt a M, Germany, which is described in this paper. In the Boliden process referred to above the sinter mix is pelletized by enveloping return fines with layers of flue dust, limestone powder, and dried galena concentrate. The roasting and agglomeration are carried out on a down-draught machine, and a slight excess of sulfur is left in the sinter product. During the smelting in the electric furnance the roast-reactions occur and a slag poor in lead and a sulfur bearing lead are formed. This latter is subsequently oxidized in a converter to obtain lead bullion and dross. The Lurgi-process achieves the maximum possible extent of the roasting reaction on the sintering machine. The wet flotation concentrates are blended with return fines (lead content 70 to 80 pet), any existing flue dusts and lead slimes—but without the

Jan 1, 1962

University of Alabama University, Alabama Mining and Metallurgical Society WILLIAM S. SPRINGER, President ROBERT E. MEAD, Secretary J. R. CUDWORTH, Faculty Sponsor MILTON H. FIES, Counselor University of Arizona Tucson, Arizona Miners' Society PERCY E. COE, President W. JAMES THOMAS, Secretary E. P. MATHEWSON, Faculty Sponsor B. R. HATCHER, Counselor

Jan 1, 1937