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By G. W. Bain

The Canadian magnesite deposits are situated in Grenville township, about 8 miles from Calumet station, on the Canadian Pacific Railway, but the poor transportation facilities have hindered their development. Magnesite was discovered in the boulders near White Rock Lake by Rev. W. P. Boshart in 1900. Observing that the material of these boulders was more glistening and whiter in color than Grenville limestone, he sent a specimen to the Department of Mines, Ottawa. An analysis showed this to be almost pure carbonate of magnesia and of commercial value if present in sufficient quantities. That summer, Boshart and McAllister found numerous outcrops, both of boulders and of bed rock, that gave promise of being magnesite of a commercial grade, and decided that this mineral deposit, if developed, would rival that at Veitsch, Austria. So they procured options on most of the land on which they knew the magnesite occurred. But the transportation difficulties were too great for the deposits to be mined at the prices offered for the product. In 1907, T. J. Waters organized the Canadian Magnesite Co. and opened a quarry on the north half of lot 18, range IX, and later on lot 15, range IX, of the same township. A lirnestone kiln was erected and the calcined rock was hauled to Calumet. In 1914, the North American Magensite Co, took over the quarries and shipped both crude and calcined magnesite until January, 1923. During 1915, S. Melkman organized the Scottish Canadian Magnesite Co. and opened a quarry on lot 15, ranges X and XI, Grenville township. A light, narrow-gage railway connects this quarry with the C. P. R. at Magnesite station. but shipments are made only during the summer months. Numerous other quarries have contributed to the output of the district but the greater part has come from the open pits of the

Jan 1, 1923

By Charles P. Berkey, Joseph H. Sinclair

This paper describes the results of a visit to the Santa Elena Peninsula, Ecuador, in January and February, 1921. On account of the complicated folding and faulting of the rocks and the peculiar occurrence of petroleum, field work was possible on only a few of the many geological problems and the results here published are only a contribution to the solution of the general problem. We know of nothing published on the geology of the district, except one account, in Spanish and German, by Theodor Wolf. The Santa Elena peninsula, the most important headland of the coast of Ecuador, lies two degrees south of the equator. The chief town, Santa Elena, is about 75 miles west of Guayaquil, the main seaport of Ecuador. The nearest developed oil field is the Zorritos of northern Peru, 120 miles south across the Gulf of Guayaquil. If newspaper reports are true, viz. that the Anglo-Ecuadorian Oilfields, Ltd., has recently brought in a 700-bbl. well on the coast of the peninsula near Ancon, the district may become important in the production of petroleum. At the time of our visit, the best way to reach the peninsula was by a steamer of the Pacific Steam Navigation Go., southward from Panama and northward from Guayaquil. About once a week, one of these steamers anchored 1/2 mile off shore near the village of Salinas, an important cable station. An alternative route, in the dry season, is by automobile from Guayaquil. In February, 1921, a railway bed had been graded part way between Guayaquil and Salinas, but this was detcriorating rapidly because of lack of funds to continue construction or to keep in repair that portion already completed.

Jan 1, 1923

By C. Erb Wuensch

The Eagle mine is situated in the Kerber Creek mining district, Bonanza, Saguache County, Colo. The climate and topography of this district arc similar to those of mining camps of the Rocky Mountain region that lie 9000 to 12,000 ft. above sea level. The district is at the northeasterly edge of the San Juan volcanic region, in which successive Tertiary lavas have built up a high plateau resting upon earlier sedimentary formations, or the underlying pre-Cambrian complex. The latter was exposed in places by erosion previous to the eruptions. In the Bonanza district, proper, only volcanic formations, andesites and latites, are found. Dr. H. B. Patton,1 in his survey of the district, discriminates between several varieties of lavas and maps these as distinct flows. They are local variations within the same lava and it is impossible to determine any distinct line of demarkation between them. Structurally they are members of the same flow. There are three general types of vein occurrences: 1. Lead-zinc-copper-silver-gold veins in well-defined fissures with a quartz gangue; ill these, the various sulfides show a zonal arrangement. 2. Massive gray copper ores high in silver, in irregular fractures in the form of small lenses and pockets, in which little or no quartz is found. 3. Manganiferous quartz veins very low in silver and gold at the surface which, with depth, become important silver veins containing subordinate amounts of gold. These veins have a quartz gangue with an abundance of Phodochrosite. The silver is associated with small amounts of sphalerite, galena, and pyrite; fluorite also is present. The Eagle is the only mine that has been developed in the manganiferous quartz type of vein. It was discovered ill the fall of 1582, but little more than assessment work was done on the property for a number of

Jan 1, 1923

By Richard B. Moore

The successful commercial production of helium during the last few years has added greatly to its scientific interest. When the quantity of an element available for experimental purposes increases within a few years by 200,000 per cent., both public and scientific interest in the element are greatly stimulated. Before the war, probably, there was not more than 10 or 15 cu. ft. of helium in existence. The larger part of this was owned by Prof. Kam-merlingh Onnes, and was used by him in connection with his low-temperature work at Leiden. By experimental work during the war and immediately afterwards, the Canadians, under Prof. J. C. McLennan, produced about 60,000 cu. ft. of helium of various grades of purity. Up to Dee. 1, 1921, when the plant at Fort Worth was shut down because of lack of funds, there had been produced during the experimental work by the U. S. Government, during the war and afterwards. 2,300,000 cu ft. of helium averaging between 92 and 95 per cent. purity. After the new appropriation was made in 1922, the production plant was again started and at the time of writing was making about 15,000 cu. ft. of helium a day. Only three of the five units were being operated because of the inability to get sufficient gas to run at full capacity. The gas supply is owned by two companies and on account of a disagreement between them, it has been possible for some time to get gas from only one. It is expected that when all five units are running steadily, the daily production will be between 30,000 and 40,000 cu. ft. of gas averaging from 93 to 95 per cent. purity. All three experimental plants used during the war have been shut down. The Air Reduction plant, using the Claude system, and the Linde

Jan 1, 1923

By Kenzo Ikeda

The Kosaka smelter is situated in the extreme northern end of Hondo (the main island of Japan) 15 mi. east of Odate, on the government railroad, to which it is connected by a private railway. It contains ten blast furnaces having a maximum daily capacity of 1600 short tons of ore, conducting a slow pyritic smelting entirely without coke or limestone, and with a charge carrying over 50 per cent. of briquetted and agglomerated fines. The matte (35 per cent. Cu) is blown in two basic-lined barrel converters, the blister copper (98.8 per cent. Cu) being cast at once into anodes. The adjoining refinery contains 600 tanks with a maximum capacity of 1300 tons of cathodes per month, the byproducts being gold, silver, and bismuth. At present, working at about half capacity, the whole plant consumes 2000 kw. of hydro-electric power, and employs 850 men and women. The ores are derived from two principal sources: The relatively new Hanaoka mine, 20 mi. from the smelter, yielding copper pyrites by underground mining, and the old Kosaka open-cut mine, 1 1/2 mi. from the smelter, yielding three different ores: (a) kuromono, a compact finegrained mixture of barite (32 per cent.), zinc blende, galena, and pyrite, the whole carrying about 2.5 per cent. Cu; (6) cupriferous pyrites with 2.4 per cent. Cu; (c) siliceous ore with 2 per cent. Cu. The Hanaoka ore now constitutes approximately one-half of the charge, almost replacing the Kosaka kuromono which was the ore chiefly smelted until 1913.'

Jan 1, 1923

By G. D. Van Ardsdale, A. W. Hudson

Heap leaching, as practiced at Rio Tinto, Spain, while one of the oldest, and probably one of the cheapest, methods of extracting copper from its ores, has not had, until recently, other than experimental application in this country. This has been due partly to a mistaken idea that the Rio Tinto ores possess some obscure and mysterious quality that renders them alone suitable to the process, partly to the fact that tests properly made meant the expenditure of large sums of money and several years time, and partly to the fact that favorable commercial and other conditions are not often found. This paper gives some views on the chemistry and mechanism of the method, an account of the preliminary experiments, and the final plans adopted for a large-scale installation of the method by the Phelps Dodge Corpn. for treating low-grade ore from Sacramento Hill, Bisbee. If results comparable with those obtained from the test heap are to be had from the large plant, this plant will operate on the lowest-grade sulfide ore now being treated commercially in the country. The method has several advantages: The installation cost is low. Interest charges must, of course, be made both on the cost of the ore delivered to the plant and for the plant itself, but as these are low, there is considerable advantage in the ability to regulate output by more or less complete shutdowns in accordance with market and other conditions. The amount of labor needed is small, and the cement copper produced may be considered nearer a finished product than the sulfide concentrates from a mill, and certainly means less installation cost per ton to put into final form by present methods. At Bisbee, conditions were favorable for a trial of the method for, in the preparation of the Sacramento Hill orebody for steam shoveling,' a large amount of material below concentrating grade had to be moved to

Jan 1, 1923

By Allan J. Clark, W. J. Sharwood

When the cyanide process came into general use, late in the nineteenth century, chlorination was quickly supplanted, but amalgamation yielded place more slowly, being still the major process at many plants, where it is followed by cyanidation of the tailings. The research work of the Denny brothers in South Africa' was probably the most important contribution of its time to the metallurgy of gold. By grading analyses of mill products, they pointed to the necessity of fine grinding to accomplish maximum extraction by cyanide. The work here described is merely an extension of the Dennys' method. Developing the fine-grinding principle, the steps of "crushing in solution" and of "all-sliming" were logical sequels, though both had been attempted early in the history of the cyanide process. Crushing in solution introduced certain complications in amalgamation; when "all-sliming " was adopted it was observed that the omission of amalgamation did not materially affect the total recovery. Thus the field of the older process has become much restricted, being at present confined to ores which, yielding freely to amalgamation, either are too low in grade to show an economic advantage in "all-sliming," or present chemical or other difficulties making the use of cyanide solution unadvisable during the crushing and preparation of the ore. South African ores are typical of the former class; Homestake ores are of both classes, but even in these instances the amalgamation work is being curtailed and greater burdens thrown on the cyanide plants. It is obvious that the time required for the dissolution of gold depends on the size of its particles. In an effort to determine the critical size, the following laboratory method was evolved. A standard amalgamation test2 had already been established, which involves the agitation of about 100 gm. of ore with 20 gm. of pure mercury, the latter then being separated and its gold contents determined. This method was applied directly to certain mill products with interesting results, a few of which may be noted here. The column "Free Gold" gives the value per ton thus amalgamated, as against the " Total Gold " determined by fire assay.

Jan 1, 1923

By Jesse O. Betterton

Recently some experimental work was conducted by the author in connection with the direct leaching of certain zinc ores with sulfuric acid with the object of subsequently recovering the zinc by electrolysis. The ores contained relatively large quantities of soluble silica, making it necessary to evolve special methods for handling the silicate solutions resulting from a leach of the ore, and also for the final removal of the silica before electrolysis. The problem was complicated by the presence of vanadium in the ore, the commercial recovery of which was very desirable. This paper, however, deals only with the silica phase of the problem. The zinc in the ore occurs in the form of a silicate (hemimorphite) and, to a very small extent, as the phosphate (hopeite) and carbonate (calamine); the vanadium is found as descloizite and vanadinite but mainly as descloizite. The experimental work was done in a plant producing 50 lb. of zinc = daily, in which the complete cycle of operations, including the final electrodeposition of the zinc, was performed continuously for one year. Some of the conclusions given here are the result of the operations in -this plant; the others are from laboratory tests inaugurated both as a guide to the plant operations and, in certain cases especially, to determine the limits in question. The practice in the 50-lb. zinc plant was to direct leach the ore crushed to pass 40 mesh, with spent electrolyte from the cells containing approximately 3 per cent. zinc and 10 per cent. free acid. This resulted in an acid leach solution and an ore residue. The acid leach solution was ' subjected to a special treatment for subsequent removal of vanadium; then the silica and vanadium in solution were precipitated together and removed as a "silica residue" by pressing through an ordinary plate and frame press. The next step was to treat the filtrate from the silica residue with zinc dust as a final purification step before it went to

Jan 1, 1923

By Arthur Crowfoot

As a result of work done, on an operating scale, in its experimental mill during 1920 and 1921, the plan here given was proposed for crushing, grinding, and concentrating low-grade sulfide ores at the Morenci Branch of the Phelps Dodge Corpn. The flow sheet is designed to recover 85 per cent. of the total copper and 95 per cent. of the sulfide copper from ore assaying 2.0 per cent. total copper and 0.3 per cent. acid soluble copper. At present, 1400 tons are treated daily under the conditions of this flow sheet, two 700-ton units having been installed in the 4000-ton concentrator. The run-of-mine ore will be reduced, in the dry-crushing plant, to —6 mesh and then delivered to the concentrator feed bins. The crushing will be done in three stages: (I) Run-of-mine ore is reduced to —2 1/2-in. in a gyratory crusher; (2) the 3-in. ore is reduced to 1/2-in. in 72 by 24 in. rolls; (3) the 1/2-in. ore is reduced to — 6-mesh size in 54 by 24 in. rolls. The 54-in. rolls will operate in closed circuit with electrically vibrated screens, "Hum-mer" type, a conveyor circuit being used to carry the roll product to the screens. Sufficient screens will be installed to pass 1.5 tons undersize per hour per square foot. Fine-ore bin chutes will be piped for compressed air, which will be used intermittently to keep the ore feed in motion if it tends to arch or pipe in the bins. The Concentrator The concentration of the copper and iron sulfides in the ore will be effected in three stages: I. By direct treatment of the —6-mesh crushing-plant product on primary tables, which will recover about 35 per cent. of the original copper content of the ore in a clean granular concentrate. The table middling will be dewatered in Dorr rake classifiers and ground to — 48-mesh in ball mills operating in closed circuit with circulating classifiers.

Jan 1, 1923

By H. E. T. Haultain

The study of rock crushing or grinding in tube-mills is difficult on account of the large size of the units employed in the field and the large number of variables entering into the problem. Three or four years ago, in the laboratory of the Department of Mining Engineering of the University of Toronto, a serious effort was commenced to study the matter under laboratory conditions. Experiments were carried on in a cylindrical tube-mill 18 in. long and 12 in. in diameter, inside measurements,

Jan 1, 1923

By F. C. Dyer, H. E. T. Haultain

There has been much written on ball-mills, but no small amount of the literature is simply the expression of individual opinion without sufficient data. This is no doubt due to the complexity and obscurity of the phenomena involved. The outstanding paper is by E. W. Davis,I superintendent, School of Mines Experiment Station, University of Minnesota, which is based on a large amount of data from careful experiments with full-size apparatus, together with some mathematical philosophy. The entire philosophy is summed up in the equations given; but the present writers question the foundation upon which this philosophy is based. Encouraged by the work of Davis, the authors of this paper continued work they had been doing along this line. At first, photographs were taken of the ends of rods in a small laboratory rod mill. The mill was tilted slightly toward the feed end and the discharge end was removed. The rods remained in the mill long enough to permit good observation and good photography. Later, a mill 24 in. inside the stepped lining was constructed. It consisted of two sheets of plate glass, 1/2 in. apart, set in a cast-iron ring. It was driven by a variable speed motor and was connected to an accurate speedometer (magneto and voltmeter type). In place of balls or rods, cast brass disks were used. These were 3/8 in. thick and varied in diameter from 2 in. down to 3/4 in. Four small projections on each side prevented the disks sticking to the glass when water was used; sometimes steel balls % in. in diameter were added to the load of disks. There was also constructed a 6-in. mill with corrugated linoleum lining and containing a dry load of various sizes of steel balls. The mills were so connected to the motor that the small mill made two revolutions to one of the large mill. Many runs were made with a great variety of materials, including crushed marble, seeds of many kinds, and many mixtures (some wet and

Jan 1, 1923

By A. B. Calhoun

These mines, which belong to the Burma Corporation, Ltd., formerly a London company now incorporated in Rangoon, Burma, are situated in the semi-independent state of Tawng-Peng, one of the small divisions comprising the northern Shan states and erroneously known as part of Upper Burma. Bawdwin is approximately 23' 6' N. latitude and 97" 20' E. longitude, 450 miles north of Rangoon, 169 miles northcast of Mandalay and 50 miles south and west of the Province of Yunnan, China (see Fig. 1). Although the mines were worked by the Chinese during the Ming Dynasty, 1412 A. D., the most extensive operations took place between 1796 and 1851, when the mines contributed largely to the silver market of China. During the reign of Tung Chik, 1868 A. D., they were abandoned partly because of the Mohammedan rebellion in Yunnan, which made life and property insecure, but largely because of the difficulty of operating the mines on account of water and poor ventilation. Since that time practically nothing was done, although the Burmese kings are said to have made several sporadic attempts to work the mines, until 1891 when Europeans were attracted by the great slag dumps (assaying about 40 per cent. lead) which the Chinese had left after extracting the silver. After a railroad and smelter had been built and 200,000 to 300,000 tons of slag smelted for lead, exploration work was started and old workings cleaned out in order to locate the remnants of the orebody that it was supposed the Chinese had left. After two years of most discouraging work, the remains of a large orebody was discovered by the Dead Chinaman Tunnel or what is the 171-ft., or No. 2 level adit. From then on the development was rapid and today the Chinaman is considered one of

Jan 1, 1923

By M. Mathieu

The Rove tunnel is the means by which the canal from Marseilles to the Rhone Riverl penetrates the hills of Nerthe, lying between Mar- seilles harbor and Lake Berre, Fig. 1. The canal will communicate at sea level with both terminals. The original project planned a lock near the village of Marignane; but observations, pursued for more than three years, showed that the water surface gradient would average only 0.015

Jan 1, 1923

By M. W. Ditto

THe conversion of the beehive coke plants, in this country, to byproduct plants has been slow, because the coal supplies were near the centers of the steel industry. With the growth of this industry, especially with its development around Chicago, it became necessary to transport large tonnages of coal from the eastern districts and then convert it into coke. The losses due to transportation costs were partly offset by the value of the byproducts recovered. To save the transportation costs, it was desirable that the coal deposits of Indiana and Illinois be utilized. This coal had always been classified as non-coking; it was also considered unsuitable for the metallurgical field because of the high ash and sulfur contents. The conditions, however, were promising enough to start experimental and development work, which crystallized in the design of the Roberts coke oven. The fundamental features of most coke ovens, with respect to the application of heating gas and the recovery of byproducts, are the same and, in the last few years, the tendency has been toward the better application of heating to the walls, higher thermal efficiency, and, by the control of heating conditions, the increase of byproduct yields. Structural features also have been improved and the use of high-grade refractory material has allowed the use of higher temperatures with the resulting higher rated capacity per oven per day. Great improvements have been made in both recuperators and regenerators, particularly with reference to the individualizing of each oven with respect to its adjacent oven, so that each can be operated as a unit if desired. The highest development of the flue type of oven has been applying individual regenerators for the recovery of waste heat. After some research work, it was found that the high-volatile Illinois coals could be utilized for coke by obtaining the better application and control of the heating conditions in the oven. Most of this high-volatile coal that had been used gave a good coke structure under certain conditions. The fact that coke was occasionally made from this coal, demonstrated the possibiity of making coke from it at all times pro-

Jan 1, 1923

By Jos. Van Ackeren

Although the Siemens regenerator principle was introduced into byproduct coke-oven design about 40 years ago, many problems of construction, and particularly of heat distribution and pressure conditions, have not been fully solved. Heat should be applied to the walls of a coke oven in such a way that the coking of the charge of coal will be completed everywhere at the same time, and at approximately the same temperature. This means that the application of heat should be uniform from the top to the bottom of the coal mass (except for the thin layer at the top that is purposely kept at a somewhat lower temperature to reduce cracking of the byproducts); also that a little more heat must be applied to the coke end than to the pusher end, because the coal mass is a little thicker at that end. As a modern coke oven is about 40 ft. long and 12 ft. high, a real problem is involved in the proper application of heat to the oven walls so as to secure quick coking time (without endangering the walls through local overheating) and produce coke made at a uniform temperature. Design of Ovens In all regenerative ovens, the points of combustion are arranged in duplicate sets, one set operating in one-half the area of the heating wall and the other set alternately operating in the other half. Because of the reversal of the flow of gases, the points of maximum and minimum temperatures are alternately located at the same places, thus preventing dangerous local overheating. The practicability of the principle of reversal of flow seems to have been recognized at an early date, and has been applied with every conceivable flue arrangement, combustion chamber, and means of air and gas supply. To summarize the various principles that have proved useful for the control of oven-wall temperature, the best practice seems to have resulted from the acceptance of the use of vertical flues; the separate supply of air and gas to the base of these flues, with the resulting flame

Jan 1, 1923

By Ralph A. Sherman, John Blizard

The experimental investigation described in this paper was conducted to determine the relative combustibilities of different samples of blastfurnace coke when burned in a current of air in a small furnace. It was undertaken in connection with the work done by Perrottl and Kinney2 on the process of combustion in the blast-furnace hearth.3 Reference is often made to the "combustibility" of coke by blastfurnace men, but exactly what is meant by the term "combustibility" is somewhat vague. Brassert,4 in a paper before a meeting of blastfurnace men in 1914, refers to combustibility as follows: The value of the coke in relation to the heat reactions in the furnace is determined not only by the carbon content but even more so by its combustibility. This, in turn, is governed by its size, the physical and chemical composition of the carbonaceous substance forming the cell walls, and the interior and surface structure of the coke pieces. These properties are, in turn, influenced by the chemical and physical properties of the coals used for coking, the density of the cake of coal in the oven, the coking time, temperature and heat regulation of the ovens, and last, but not least, by the method of quenching, screening and handling the finished coke. What principally concerns the blast furnace is the rate of progression of the combustion, which depends not so much on the chemical analysis as on the physical qualities of the coke. It is this rate of progression that we term combustibility, which is the speed at which the carbon molecules in the coke combine with oxygen under given conditions. In general, what the blast furnace requires is a coke which will burn rapidly and with intensity at the tuyeres, yet which has a sufficiently strong structure and hard surface to resist abrasion and solution by CO2 in the upper part of the stack.

Jan 1, 1923

By G. St. J. Perrot, S. P. Kinney

Consumers of metallurgical coke are agreed that the quality of their fuel plays an important part in the performance of the furnace. Less unanimous agreement is evident when the properties of a desirable coke are discussed or when consideration is made of suitable testing methods by which the worth of a coke for metallurgical purposes may be determined quantitatively. Much has been written of the importance of combustibility as a factor in determining the performance of coke in the blast furnace and it is generally believed that cokes vary widely in this respect. The usual operating data of the furnace afford little evidence as to the combustibility of the fuel. As a part of a general investigation of the properties of metallurgical coke, the Bureau of Mines has explored the hearth combustion zone in a number of blast furnaces by means of water-cooled gas sampling tubes driven into the hearth through the tuyeres. The attempt has been made to determine the difference in combustibility of cokes varying widely in physical properties. Data have also been obtained in an experimental combustion furnace at the Bureau laboratories in Pittsburgh. This paper gives the results of experiments, made at eleven blast furnaces, in which the composition of the gases was determined at various points in the hearth in the plane of the tuyeres. Further work is in progress in which the zone above the tuyeres will be explored. Factors Affecting Coke Combustibility Brassertl has defined combustibility as the rate of progression of combustion, or, in other words, the speed at which the carbon molecules

Jan 1, 1923

By S. P. Kinney

Sulfur has been one of the most troublesome elements encountered since the earliest days of iron smelting, and this problem will become of increasing importance as the higher sulfur coke is used, because of the rapidly diminishing supply of low-sulfur coals suitable for the manufacture of metallurgical coke. The deleterious effect of sulfur in iron and steel is well known; therefore, except in special cases, it is desirable to make the iron produced by the blast furnace as low in sulfur as possible. Sulfur enters the furnace in all of the raw materials, but principally in the coke. It comes out of the blast furnace distributed between the iron and the slag; very little passes out of the furnace with the gases. The mechanism of the reactions that cause the sulfur of the coke to pass into the slag and iron has not been studied to any great extent, although the character of the end products, and the conditions controlling the distribution of the sulfur between the iron and slag, are fairly well known. In the iron, the sulfur occurs as ferrous sulfide; or manganese sulfide in case of iron high in manganese. In the slag, the sulfur is present mostly as calcium sulfide. The presence of free sulfur1 and sulfates in blast-furnace slags is to be accounted for on the basis of oxidation of the calcium sulfide after the slag is tapped. The generally accepted reaction for the relation between the ferrous sulfide in the iron and in the coke, and the calcium sulfide in the slag, is that given by Howe? FeS + CaO + CDFe + CaS + CO

Jan 1, 1923

By Joseph F. Oesterele, Richard S. McCaffery

This investigation was undertaken to determine the quality of different iron blast-furnace slags as desulfurizing agents, and the possibility of using, in the blast furnace, materials of higher sulfur content than is now common. The solution of the problem would render possible the use of higher sulfur iron ores for pig iron and cokes containing more sulfur than now employed, and the economic manufacture of irons and steels of lower sulfur content. It is generally assumed that the sulfur that reaches the lower part of the iron blast furnace is removed from the pig iron by combining with lime to form calcium sulfide. Most of the sulfur introduced with the charge is in combination with iron or manganese or, perhaps in the coke, as a carbon compound.1 The iron and manganese sulfides are more soluble in the iron than is calcium sulfide, and as the latter is generally said to to be more soluble in the slag than the iron and manganese sulfides, this greater solubility of calcium sulfide in the slag tends to free the metal from sulfur. This greater solubility of the calcium sulfide is used not only to reduce the sulfur contents of the pig iron in the blast furnace, but by this means also low-sulfur steels are produced in the open-hearth and electric furnaces. The reaction2 FeS + CaO + CO = Fe + CaS + CO2 requires a high temperature and strongly reducing conditions in the presence of lime. The constitution of blast-furnace slags, first studied scientifically by Vogt13 indicated certain ranges of composition of the slag in which a

Jan 1, 1923

By G. P. Pilling

The use of magnetic ore in the blast furnace is a subject of increasing importance. The end of the deposits of lake ore is in sight, although not imminent, and unless some new field is discovered, the use of magnetites will be an increasing factor in the iron production of the country. Enormous deposits of magnetic ore in the country that have been practically untouched contain ore of sufficient richness and purity to prolong materially our ore supply without extensive importation of foreign ores. Magnetic ores can be divided into three general classes: Fine un-sintered, or raw ore; raw lumpy ore; fine ore that has been sintered. Fine Unsintered Ore Typical analysis and screen test of this class of ore from the Adirondack region is as follows: Per Cent. SCREEN TeST pER Cent. Fe.................... 58.00 Remaining on 4 mesh........... 7.50 SiOz.................. 7.50 Through 4 on 8 mesh........ 13.75 Al,O,................. 1.60 8 on 10 meah........ 3.75 Baseg................. 2.77 10 on 20 mesh........ 22.50 P.................... 1.00 20 on 40 mesh........ 27.50 Mn................... 0.13 40 on 60 mesh........ 2.50 S..................... Trace 60 on 80 mesh........ 10.00 80 on 100 mesh........ 2.50 100 mesh........ 10.00 100.00

Jan 1, 1923