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By Nicholas J. Grant, Oliver Preston

PUBLICATION of data by Irmann' indicating outstanding thermal stability and elevated-temperature strength properties in a sintered aluminum powder product (SAP) stimulated interest in the strengthening of metals by means of a highly dispersed oxide phase. This material was produced by compacting, sintering and extruding a fine flake aluminum powder, resulting in a dense product containing 10 to 16 volume pct alumina, which arises from the thin oxide film on the starting powder, highly dispersed in a matrix of pure aluminum. The method of producing SAP, and the stable high-temperature properties obtained from it, introduced the possibility of developing superior materials for high-temperature service by powder metallurgy methods. The composition of the materials of this type which are presently under consideration is predominantly metallic, containing from less than 1 pct up to about 20 volume pct of the oxide. As a result, other properties of these materials, such as electrical and thermal conductivity, thermal shock resistance, and to a certain extent ductility, notch sensitivity, machinability and hot working properties, are also predominantly metallic. These alloys do not constitute, therefore, a completely new class of materials as do the cermets, but rather present a new means of modifying the properties of a metallic matrix. In this sense, SAP type alloys are more closely analogous to the alloy aging systems. While the strength properties realized in the oxide-pure metal systems at the lower temperatures are substantially less than the best attainable in the aging systems, the thermal instabilities in the latter make them of less value for use at temperature much above their optimum aging temperatures, as indicated in Fig. 1. While the means of obtaining an oxide dispersion are more difficult and costly than conventional aging processes, in the light of the as yet limited understanding of such systems there is considerable promise of improved high temperature properties arising from the thermal stability of these structures. Since the development of SAP in 1946, its properties and the properties of a series of similarAl,O,-A1 products containing from about 1 to 20 pct aluminum oxide have been fairly extensively studied and reported.'- "articular attention has been given to

Jan 1, 1958

By L. Bombardieri, H. F. Mills

Iron King mine, producing gold-silver-lead-zinc-ore, is 10 miles east of Prescott, Ariz. At present the 1806 level is being developed. The echelon pat tern of ore deposit continues at depth but is less pronounced, and the orebody is now more or less continuous over 2600 ft of strike length, with widths 5 ft to as high as 26 ft where overlapping of lenses occurs in one section of the mine. Better continuity of ore at depth has encouraged systematic mining and improved plant facilities to offset higher operating costs. With favorable persistence of ore at depth, the three-compartment No. 6 hoisting shaft appeared inadequate to handle increased tonnage from greater depths, and in 1952, a four-compartment steel and concrete shaft was started. This is now at a depth of 1906 ft, and shaft pockets and loading cartridges are being installed on the 1806 level. Ore from the new shaft, hoisted in 5 ½ -ton Jeto bottom dump skips, is dumped into a 500-ton shaft bin, from which it is fed to a 16x30-in. gyratory crusher and conveyed 415 ft through an undersurface concrete gallery to the secondary crusher and conveying system supplying the 900-ton mill. Lead, zinc, and pyrite concentrates are floated, and the flotation tailings are cyanided after desliming. Waste from the new shaft is dumped into a 250-ton surge bin, fed from this bin to a 24-in. conveyor, and discharged into the glory hole above the block cave. The new shaft is served by a double-drum 400-hp hoist used for ore and waste; the No. 6 shaft conveys men and material.

Jul 1, 1956

By L. A. Delano

During 1916, the St. Joseph Lead Co. milled 2,505,670 tons of ore. This is a daily operating average of 7855 tons. The economic concentration of such a large tonnage necessarily requires a plant equipped with modern machinery and operated efficiently. The milling process used in producing this output has been a gradual development through the long period of years that the company has been in existence. HISTORICAL SUMMARY OF MILLING PRACTICE Ore was first mined at Bonne Terre, Mo., in 1864. The early milling practice was crude. It consisted of breaking the ore into small pieces by hand, pulverizing in a Blake crusher and a pair of Cornish rolls to approximately minus 6 mm., and jigging in the old-fashioned hand jigs. Only the coarse heavy galena was saved, the gangue being shoveled off and discarded. The output for the first year under this method was only 240 tons of pig lead. The first mill of the company was a combination of extensions and additions built from time to time as the mines developed. It was a museum of early milling machinery, such as jigs, percussion tables, buddies, dolly tubs, log washers, etc., following no definite flow sheet. The chats, or tailings, were hauled to waste dumps by mule and ox teams, as much as 500 tons per day being moved in this manner, until a railroad was built in 1880. This mill was burned in 1883, at which time the annual output had reached approximately 7300 tons of pig lead. A new mill of 500 tons capacity was designed and built by C. B. Parsons. This mill has been admirably described by H. S. Munroe.' The ore was crushed in Blake crushers, Cornish rolls, through 6-mm. screens, and treated by jigs of the Parsons type, percussion tables and log washers, or trunking machines, for cleaning the galena. The middlings were recrushed in rolls and treated on Harz jigs. The mill was operated without change, except later installing Gates crushers and en-

Jan 1, 1918

By Leonard Grimshaw

DIFFUSION Of carbon from gases into iron has been the object of much research, because of its long recognized importance in carburizing processes, but the direct diffusion of carbon from steel into iron has received relatively little attention. Because of the increasing use of clad metals in industry, direct diffusion from steel into iron is becoming of practical importance. Therefore, the object of this investigation was to study the diffusion of carbon from several different kinds of steel directly into relatively pure iron, realizing that elements in the steels other than carbon might influence the rate of carbon diffusion. The previous work on diffusion of carbon into iron has been consider-able, and includes many reports in foreign languages. This work has recently been well reviewed by Mehl1 and Krynitsky,2 and the reader is referred to their excellent bibliographies.

Jan 1, 1937

By A. W. Gauger

Coal as mined contains varying quantities of inorganic components (mineral matter) which, on combustion, produce the residue known as ash. It has long been realized that the weight of this residue does not equal the weight of the mineral matter originally in the coal, because of the chemical changes that take place during the ashing process. Moreover, the ordinary ash analysis presents the ash as a mixture of oxides with little or no clue to the identity of the mineral species originally present in the coal. It is only for pyritic sulfur that an attempt is made to determine a mineral component. So far as the authors are aware, there are practically no experimental data in the literature as to the percentages of the different mineral species present in coal. The oft-quoted table of Marson and Cobb1 cannot rightly be considered as data, since the authors merely state that the table "shows the probable forms in which these ash constituents exist in coal." In view of the attempts to calculate analyses and determinations of the calorific equivalent of coal to the mineral-matter-free basis, as well as the importance of the mineral species in clinkering, combustion and coking, it seems strange that so few data are available on this phase of coal technology. For example, numerous empirical corrections for use in the scientific classification of coal have been suggested to compensate for the losses from the mineral matter during ignition of the coal to ash. These include multiplication of the ash percentage by various factors such as l 1/10 and 156, the use of the Parr unit coal formula (1.08 A + 22/40 S, where S = total sulfur) as well as certain modifications of the Parr formula that attempt to correct for the organic sulfur present in the coal. Although many experiments have been made to determine the loss on ignition of the clays associated with the coal measures and correction

Jan 1, 1934

By Robert E. Wilson

TO show the vital importance of our future oil supplies to our economy, I will merely point out that this country, with something like 15 per cent of the world's land area and something like 7 per cent of the world's population normally uses between 65 and 70 per cent of the motor vehicles of the world, between 65 and 70 per cent of the world's rubber, and produces and refines between 65 and 70 per cent of the world's petroleum. Despite all the arguments as to which is cause and which is effect, all will agree that all three of these important industries are mutually dependent. If we had not been able to produce 65 or 70 per cent of the world's oil in this country we would not have the degree of mechanization, or the efficiency of transportation, or the high industrialization that we enjoy. Furthermore, the fact that oil has been plentiful and cheap has had a tremendous effect on all the things in which we as engineers are interested. For that reason, if for no other, it is well to look ahead to see what the outlook is for the near and far-distant future.

Jan 1, 1945

By Henry A. Wentworth

(New York Meeting, February, 1912.) ELECTROSTATIC separation of ores in its present form is generally known as the Huff' process from the name of Charley H. Huff, of Boston, Mass., through whose constant and persistent labors (with the invention of Clinton E. Dolbear as a basis) the successful commercial process embracing separative machinery and the various electrifying devices has been developed step by step, and the finances for the long period of development provided, and the method finally established and recognized throughout the world as an important and successful addition to the ore-dressing department of metallurgy. The permanent field-success of electrostatic separation began in 1908 with a 20-ton Huff experimental mill built specially for the purpose by the American Zinc, Lead.& Smelting Co., in Platteville, Wis., a plant which was a success from the start and was gradually increased in capacity as the market-conditions warranted to 100 tons of concentrates per day. Much credit is due to the above-mentioned company for its initial venture, and for its assistance in applying the process to field use. Prior to 1908 electrostatic separators had been installed and operated (but for a comparatively short time, however) in a number of places; some under the patents of Mr. Dolbear by himself and associates, and some under the patents of Lucien I. Blake and Lawrence N. Morscher, by W. G. Swart, mining engineer, of Denver, who has always been a courageous advocate of electrostatic separation. Due to the difficulties experienced in the generation of the electrical charges, the primitiveness of the separators, the wooden construction instead of iron as at present, the lack of control of the electrical fields and other difficulties overcome by the later inventions of the Huff Electrostatic Separator Co.,

Jun 1, 1912

By T. M. Chance

THe necessity of adopting improved methods of bituminous-coal cleaning at many collieries is recognized by all familiar with the requirements of the various bituminous markets, the approaching exhaustion of the relatively small areas of inherently high-grade coal, and the difficulties encountered in maintaining a uniform grade of shipped coal. Many coals lose much of their market value by degradation in size, and, in any event, if the cleaning .process is to contemplate wetting the coal the minimum amount of fines should be produced to prevent sludge losses and the difficulties resulting from the moisture content of the wet fines. Hence, it is desirable to bypass the finer coal dry, if this can be done without too great increase in impurities in the final product. This procedure has given excellent results at washeries at Steelton and in the Birmingham district. If the raw fines are too impure to permit admixture with the washed coal, it may be desirable to use dry methods, such as the pneumatic table, rather than to wet an excessive amount of such fine material. The bituminous-coal cleaning plant should therefore: (1) Treat the raw coal in as large lump as possible and without sizing; (2) eliminate degradation of pure coal in the washing process; (3) avoid wetting the fines, if possible; (4) produce washed coal practically free from mechanically included impurities; (5) crush the refuse and retreat, if much intergrown coal and refuse are present, rather than precrush the entire feed. Ordinary gravimetric methods of concentration, whether hydraulic or pneumatic, will not meet all of these requirements, as the large lump includes those sizes now sent to the picking tables, if it is desired to treat material of such size. I do not know of any prior commercial method, based on the falling velocity, coefficient of friction or resistance to fracture, that permits the production of clean coal varying in size from, say, 10-in. lump to 1/16-in. mesh. The effect on the specific gravity of any given coal as a result of variations in ash and pyritic sulfur content is well known; in fact, it has been made use of in commercially ascertaining ash content by deter-

Jan 1, 1924

By F. C. NINNIS

AT THE Belmont mill, the pregnant solution is de¬livered to a 30 by 10-ft. tank, from which it is pumped to three Merrill clarifying presses of the sluice-bar type, whence it flows through the meter to the precipitation supply tank. ?The solution for precipitation is pumped by a 10 by 10-in. Aldrich triplex pump and delivered through a 6-in. line to the forty-eight triangular frame Merrill presses at the head of the mill. Zinc dust is fed in at the pump suction and the solution precipitated to 0.01 to 0.02 oz. silver. The solution flowing from the presses is tested at 1-hr. intervals with sodium sulfide. All solution is precipitated to the same value; 4J tons of solution are precipitated for each ton of ore treated and the presses are cleaned every 15 days. The product from the presses' is dumped into steel cars having steam-jacketed bottoms, in which they are dried until the moisture content is between 15 and 17 per cent. when 8. per cent. of soda and .3 per cent. of borax are added and the contents thoroughly mixed. In the case of special clean-ups, the precipitate is tested and the flux is added to meet the requirements.

Jan 1, 1921

By Robert J. Ramsey, Robert D. Jeppson

Introduction The Utah Copper Division of Kennecott Copper Corporation will present its contribution to the A. M. Gaudin Flotation Symposium in four parts. The first two segments will discuss briefly the historical development of the Utah Copper Division and describe the ore body. The third segment will review the facilities and operating procedures for the Utah Copper division's newest plant, the Bonneville Concentrator. The fourth segment describes flotation process circuitry, including molybdenite recovery procedures, at the Arthur and Magna concentrators. (The Bonneville concentrator is a crushing and grinding unit and does not have its own flotation plant). I. Historical Development Progress through change has been the key to the successful operation of the Utah Copper Division for more than 70 years. As the original low-grade copper operation, the pressures to remain competitive and profitable have dictated the timing and extent of each change. Much of the change has occurred through technological and economic advancements in the area of ore concentration. The Utah Copper Division of Kennecott Copper Corporation had its origin June 4, 1903, when the Utah Copper Company was organized. The initial mill of 300 tpd (tons per day) capacity was built at Copperton a few miles below the present Bingham Canyon mine site. From 1904 to 1910, the Copperton mill was expanded to a capacity of 1,000 tpd, and all of the known apparatus for gravity concentration were evaluated.

Jan 1, 1976

By Petri Bryk, John Ryselin, Rolf Malmstrom, Jorma Honkasalo

THE theoretical possibilities for the realization of flash smelting have been known for a long time. Calculations concerning the same can be found in previously published literature,1 and suggestions for its accomplishment in practice are described in patents on the subject. Likewise a smaller number of practical test results are known. Autogenous smelting is closely connected with flash smelting, since in flash smelting techniques it is possible to use the heat evolved by oxidation of the sulfides for the smelting process to such an extent that under certain conditions smelting can proceed without additional fuel. However, in the event that insufficient heat is available from the oxidation of the sulfide concentrates, then additional heat may be supplied by preheating the air, using oxygen enriched air, on burning additional fuel with the concentrate.

Jan 6, 1958

By Fred E. Carter

AT first sight, the platinum group of metals seem of little import to we, the people," although actually the life of the common man is much influenced by them; this influence is usually indirect, hence his lack of knowledge and interest. He generally associates platinum with high-class jewelry and lets it go at that. Were 11e to go deeper, he would find that platinum metals play their part in his food supply, his clothing. his health, his transportation, his home. And the technical man is in general only slightly better informed, for he usually does not realize how necessary platinum metals are in the manufacture of the many construction materials and chemical products with which he is acquainted. Certainly, our daily lives would be profoundly affected if platinum were suddenly to become unavailable.

Jan 1, 1944

By Thunaes, Arvid

Pilot plant test work in 1942 and 1943 showed that by a combination of desliming, fine-size classification, and Sullivan deck concentration it is possible to recover heavy minerals such as cassiterite at least as fine as 10 microns in size. This appreciable improvement in gravity concentration practice has been substantiated by several full-sized plants.

Jan 1, 1950

By Theodore Simons

During the last half century a great amount of ingenuity and energy has been devoted to the invention of appliances for the recovery of valuable minerals from very fine sands and slimes. The reason for this is that in almost every dressing plant the greatest losses of values, considered relatively as well as absolutely, occur in the treatment of slimes. The natural aim of mill operators to minimize these losses has in recent years received another impetus, from the fact that the gradually diminishing occurrence of high-grade ores makes a more intense recovery of values from existing resources absolutely necessary for profitable operation. The results of such attempts are found in the appearance of a great number of machines and appliances which make it profitable to-day to rework old clumps, containing the tailings of older processes. A constantly growing difficulty of obtaining skilled labor and the increase in wages made it prohibitory to employ those types of machines which required much attention, and in the middle of the ninieteenth century the various concentrating tables of the percussion type, which automatically discharge their products, began to take the place of the intermittent tables, on which, when the bed had reached a certain thickness, operation had to be stopped in order to allow the skimming by hand of the separate layers formed on the table. In the course of years, two general types of machines evolved for the treatment of fine sands arid slimes: The so-called concentrators and vanners, on the one hand, and the buddles and round tables on the other.' For the profitable treatment of the very finest slimes, round tables have as a last resort proved more satisfactory and econoillical than other machines. In recent years they have agaiu become the subject of close inves-

Jan 1, 1914

By Clarence O. Babcock

A simple robust rock bolt load gage was tested in the laboratory to define its characteristic behavior and to obtain a measure of its performance. The gage uses three vibrating wire sensors crimped 120 (2.09 rad) apart into the outer surface of a hollow cylindrical steel, tube. The resonant frequencies of the wires change as the gage is loaded. This change is read with a Creare meter developed for the Bureau of Mines on a previous contract to read borehole stress gage results. A statistical analysis of data obtained defines the variables influencing results. A U.S. patent (3,914,992) assigned by the author to the Dept. of the Interior was granted Oct. 28, 1975.

Jan 1, 1978

By Thomas F. Evans

This study shows that statistics on the rate of coal dumping can measure the effectiveness of freeze conditioning agents. The use of appropriately indexed dumping intervals or interval logrithms provides a more sensitive indication of unloading problems than does a simple calculation of average loading time because of the abnormal form of the distribution of dumping intervals. The paper presents information that indicates suitably applied coal freeze conditioning agents are effective in reducing coal handling problems attributable to frozen coal. Graphs and figures show that the severity of coal freezing potential due to subfreezing temperatures is proportional to the exposure time and average departure of the temperature from the freezing point.

Jan 1, 1982

By R. E. Hudrlik, G. W. Preckshot

The diffusion coefficients of silver from solid silver in molten lead were measured to within ± 0.8 pet in a columnar type diffusion cell ower, the temperature range of 326° to 530°C. Fick's law describes the process up to 530°C where the laminar mechanism appareltly breaks down. These is negligible resistance at the interface as shown by mathematical analyses. The diffusion coefficients are found concentration independent. IT would seem that diffusion in liquid metals would be free of such effects as molecular structure, dissociation. polarization. and compound formation. This view was taken by Gorman and preckshot in their study of diffusion of copper from solid copper into molten lead. They reported diffusion coefficients which were independent of the concentration over the range of 478° to 750°C. They found that the Stokes-Einstein equation with constant radius of the diffusing specie represented the diffusion data better than Eyring's rate theory equation and Sheibel's correlation. The radius of diffusion was found to be that of the doubly charged copper. There appeared to be no resistance across the solid-liquid boundary. In the present work the diffusion coefficients for silver in liquid lead were measured over a range of temperatures of 350° to 505°C. The solubility of silver in lead over the range of 303° to 630°C was also obtained. These results are compared with calculated or correlated values or with data in the literature. EXPERIMENTAL Procedure—The experimental equipment techniques and procedures were those reported in detail by Gorman and preckshot9 and will not be repeated here. Measured values of WT, Co, A. L were obtained for various diffusion times and the diffusion coefficient was computed for the case of no resistance at the interface9, 11 by: WT/CoAL = 1- 8/p2 n=1 1/(2n - 1)2 exp[-(2n - 1)2p2 Dt/4L2] [1] or where there was resistance at the interface by: WT = 1- ?n=1 2h2/ap2L [sxp [-Dan2t]/[(h2 + an2) L + h] The roots an are those of the transcendental equation3 tan (an L) = Iz/cun. The diffusion coefficient is that defined by Hartley and Crank.7 The total silver in the lead cylinder and equilibrium slug was determined by a cupellation technique' with proper correction for losses. Analysis of known samples showed that this method is surprisingly accurate. The amount of silver in the lead adhering to the silver cylinder was obtained in the same fashion as shown by Gorman and preckshot.9 The small errors involved in this determination are not critical since the silver in this adhering lead layer is only 3 to 15 pet of the total diffused. Materials—Electrolytic silver containing 99.9+ pet Ag obtained from General Refineries of Minneapolis, Minn. was used for all but runs 7 and 8. For the balance of the runs this silver was reduced with hydrogen at 1100°C and its oxygen content was found to be about 0.017 pet. For the runs. 7 and 8, phosphorous-reduced silver of the same purity was obtained from Handy and Harman Co. of Chicago, Ill. The densities of the phosphorus-reduced silver and the hydrogen-reduced electrolytic silver were 10.484 and 10.487 g per cm3, respectively. These values agree with those reported for pure silver. Silver which was reduced at 900 C had an average density of 9.998 g per cm3, indicating porosity. This silver was used for a number of runs which were not tabulated in Table I. These are indicated by crosses on Fig. 2. The 99.999 pet Pb was obtained from the National Lead Co. Research Laboratory of Brooklyn, New York. DISCUSSION OF RESULTS The diffusion and solubility results are reported in Table I for eleven runs using either phosphorus-reduced electrolytic silver or hydrogen-reduced silver at 1100° C. The solubility data shown in Fig. 1 show the excellent agreement with that reported by Heycock and Neville.8 The data of Friedrichs5 apparently are in error. The experimental solubility data of this work are reported to 0.3 pet. The experimental diffusion coefficients computed from Eq. [1] are reported within 1.2 pet of the mean and are plotted in Fig. 2. These are expressed within +0.8 pet of the experimental values over the entire temperature range by: D= 8.26 x 10 -5 e-1925/RT . [3] There appears to be little difference due to the

Jan 1, 1961

By H. Tabachnick, N. Hedley

The extraction of gold and silver from ores with alkaline cyanide solutions is well known. Cyanide solutions are also good solvents for many base metal minerals, particularly most of the copper minerals. Leaver and Woolf,1 who determined the solubilities of copper minerals in 0.10 pct sodium cyanide solution, showed that malachite, cuprite, chalcocite, and bornite are readily soluble in cyanide; least soluble are chalcopyrite and chrysocolla. The Mining Chemicals Laboratory of American Cyanamid Co. has studied the possibilities of using cyanide solutions for leaching base metals. In this article the authors present the results of their experiments with cyanide leaching to reduce the copper content of a zinc concentrate from 0.050 pct to the desired specification of less than 0.02 pct. Zinc Ore Sample: The sample of commercial zinc ore used in this investigation to produce zinc concentrates was ball mill feed ranging from 1/2 in. to fines. The ore was crushed to —10 mesh and riffled into 600-g charges for zinc flotation. Analysis revealed 4.54 pct Zn and 0.005 pct Cu. Examined microscopically, the ore consisted essentially of cream-colored sphalerite and carbonate gangue. It contained a small amount of pyrite, partly tarnished. The only copper mineral identified was chalcopyrite, mostly smaller than 20µ, some of which occurred as attachments on sphalerite. Flotation Procedure: Batches of zinc concentrates were made as required according to the following general procedure. A 600-g charge of —10 mesh ore was ground for 4 min at 60 pct solids in a laboratory steel rod mill. Screen sizing showed it to be all —35 mesh with 42 pct —200 mesh. The pulp was transferred to a Fagergren laboratory flotation machine, diluted to 25 pct solids and conditioned for 5 min with 0.6 lb per ton CuSO4, . -5H2O. It was then conditioned with 0.10 lb per ton Sodium Aerofloat promoter and floated for 5 min with pine oil as frother and creosote as froth stiff-ener. The pH during flotation was 8.5. A typical analysis of zinc concentrate subjected to leaching is 44.41 pct Zn and 0.051 pct Cu. The ratio of concentration was close to 10 to 1. Leaching Procedure: Aero Brand cyanide, technical calcium cyanide analyzing 49.5 pet NaCN equivalent, was used for all cyanide leaching tests. For each test three rougher concentrates floated as described were combined (total weight close to 180 g) and the water content adjusted to 150 ml. The percent solids in the pulp was thus 55. After a weighed amount of cyanide flakes had been added, the pulp was agitated on rolls for a given period, then filtered and washed. Filtrate and washed concentrate were assayed for copper content. The procedure for ammonia leaching was similar to the cyanide process. A series of tests was run in which the effects of ammonia as a solvent for copper in the zinc rougher concentrate were investigated. In each test 180 g of zinc rougher concentrate was leached with 150 ml of ammonia solution. Various strengths of ammonia and times of leaching were investigated. The results are shown in Table I. It is apparent that ammonia was a poor solvent for copper in the zinc concentrate; 60 min of agitation with a relatively strong NH:, solution extracted only 2 pct of the copper. CYANIDE LEACHING TESTS Effect of Solution Strength: The solvent effects of 0.125, 0.187, and 0.250 pct NaCN equivalent solutions were compared for leaching periods of 15 min. Residue and filtrate were assayed for copper content. The copper content of the filtrate plus wash was calculated back to the original volume of cyanide solution (150 ml). The effects of cyanide strength are given in Table II, It will be noted that 15 min leaching with 0.25 pct NaCN equivalent solution extracted 74.6 pct of the copper in the zinc concentrate. The leached product assayed 0.013 pct Cu, appreciably below the required specification of 0.020 pct. Effect of Leaching Time: Portions of zinc concentrate were leached with 0.25 pct NaCN equivalent

Jan 1, 1961

By CHRIS A. HARTNAGEL

For the second consecutive year, the production of crude petroleum in New York has fallen below the 5,000,000-bbl. Mark that had prevailed previously since 1937. In 1945, the output totaled 4,658,000 bbl., a decrease of slightly more, than one per cent from the 1944 production. This total represents 2 per cent of the stripper-well production of the United States, which in turn is I 3 per cent of the total annual yield of the nation's petroleum.

Jan 1, 1946

By W. E. C. Eustis

1st. The fine iron ore is mixed with a sufficient proportion of fine coking coal, and is coked in any of the ordinary methods for making coke. The effect of this is to convert the iron oxide into sponge in such a shape that the usual trouble of oxidation is avoided. 2d. The resulting mixture of coke and sponge is melted down ill an ordinary cupola, the coke furnishing more than sufficient fuel for that purpose. I may briefly point out here that in the first place the iron is sponged, and this sponge, as is well known, contains no phosphoras or silicon when the reduction takes place at a low temperature; and in the second place this sponge is melted down in a cupola in which there is no zone of reduction for phosphorus or silicon, and that con . sequently the resulting iron should be free from both of these impurities. 3d. The liquid sponge, now carburized, is run into an open-hearth furnace, where the requisite degree of hardness or softness is obtained by adding some of the fine ore, and the usual dose of ferromanganese, I have merely given above the rough outlines of the prows, bat it is my intention, as early as possible, to bring the process more fully before the Institute.

Jan 1, 1881