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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 T. F. Witherbee

The following calculation of heat requirement covers the working of the furnace from January 25th to February 14th, inclusive. A short time previous to the first date the furnace had been working rather badly, so that 300 lbs. of ore and 500 lbs. of limestone were taken off. Within the time mentioned the quality of the iron passed

By G. H. Gilman

The campaign now being waged to improve the quality of the rock-drill bit is the natural outcome of the scientific development of the drilling machine during the past twenty years. In this development there has been a great increase of power output per unit of weight of machine. But the development of the rock drilling engine has been seriously hampered by the need of a drill bit that would withstand, without undue wear or breakage, the heavy rapid blows of the improved pneumatic hammer. This problem was difficult to solve because an increase in the weight or section of the drill-steel bar resulted in decreased cutting speed and higher operating costs of both the bit and its actuating engine. The solution, therefore, is to make the bit of steel that, for a given weight or section, will be more shock and wear resisting. In the attempt to meet these requirements, many of the drill-steel manufacturers alloyed the steel with vanadium, chromium, nickel, etc. on the assumption that such elements would improve the rock-drill bit but due to the sensitiveness of such steel to forging and heat treatment, higher cost, and lack of continuity of effort on the part of the steel manufacturer to perfect this product, but little progress has been made in the adoption of these steels and the so-called straight carbon steel is today recognized as the standard material for all rock-drill bits. The first step in improving drill-bit efficiency is the determination of the cause of breakage and wear. For this reason, in 1917, certain rock-drill manufacturers, in conjunction with some of the larger mining interests, conducted exhaustive tests in both the field and the laboratory. It was then found that, in the main, the cause of drill-steel troubles is not attributable to the duty imposed upon it by the normal operation of the drilling engine, but to improper forging and heat treating of the bit and to the subjecting of the bit to improper use. The fact that it is possible to minimize the troubles attributable to improper heat treatment of the drill steel has been demonstrated and it is recommended that such knowledge be made available by teaching those interested in the production, use, and up-keep of drill bits the principles on which success in this direction may be achieved. Unfortunately, but little attention has been given to the education of the drill smith. But a few years ago, the instructions given the man

Jan 1, 1922

By F. E. Bash

About three years ago, the writer presented a paper1 on the rate of heating and cooling of a 24-in. round ingot. The present paper deals with work done on larger ingots at the plant of the Allis Chalmers Mfg. Co., Milwaukee, Wis., in August, 1919. Sufficient data on ingots of all sizes were desired to enable one to calculate in advance the rate at which any ingot could be raised to a forging temperature without injury to the steel, determining also the temperature gradient between different sections of the ingot and the length of time necessary to heat to the center. This has now been experimentally determined for 24-in. and 45-in. ignots by the writer and for a 18-in. cube by E. F. Law.2 J. F. Harper has also made tests on heat-treating temperatures on 30-in, ingots; which data are included in this paper. No attempt is made here to work out a formula for the heat penetration of steel, for too many factors are involved concerning which we have not sufficient information, as for instance, the specific heat of steel up to 2300' F. It is hoped that eventually a set of formulas will be developed so that one may calculate maximum allowable temperature differences within an ingot for any particular kind of steel and the rate of absorption of heat and temperature gradient within the ingot for a definite temperature head. EXPEEIMENTAL INGOT The ingot tested was 45 in. in diameter, having been forged into the shape shown in Fig. 1 from 54 in. diameter. Seven -in. holes were

Jan 1, 1923

By F. E. Bash

About three years ago, the writer presented a paper1 on the rate of heating and cooling of a 24-in. round ingot. The present paper deals with work done on larger ingots at the plant of the Allis Chalmers Mfg. Co., Milwaukee, Wis., in August, 1919. Sufficient data on ingots of all sizes were desired to enable one to calculate in advance the rate at which any ingot could be raised to a forging temperature without injury to the steel, determining also the temperature gradient between different sections of the ingot and the length of time necessary to heat to the center. This has now been experimentally determined for 24-in. and 45-in. ignots by the writer and for a 18-in. cube by E. F. Law.2 J. F. Harper has also made tests on heat-treating temperatures on 30-in, ingots; which data are included in this paper. No attempt is made here to work out a formula for the heat penetration of steel, for too many factors are involved concerning which we have not sufficient information, as for instance, the specific heat of steel up to 2300' F. It is hoped that eventually a set of formulas will be developed so that one may calculate maximum allowable temperature differences within an ingot for any particular kind of steel and the rate of absorption of heat and temperature gradient within the ingot for a definite temperature head. EXPEEIMENTAL INGOT The ingot tested was 45 in. in diameter, having been forged into the shape shown in Fig. 1 from 54 in. diameter. Seven -in. holes were

Jan 1, 1923

By C. R. Wohrman

A melt of pure electrolytic iron with about 0.4 per cent. sulfur and 7 per cent. manganese was prepared in connection with a study of inclusions in iron. The alloy darkened rapidly when etched with a 10 per cent, alcoholic solution of nitric acid, and microscopic examination revealed the presence of a semimartensitic, semi-Widmanstättian structure (Figs. 1 and 2) closely resembling the structures found by Benedicks in his artificially prepared "meteoric iron."' The structures found by Benedicks were readily explained by the presence in his alloy of the two well-known iron-nickel solid solutions, taenite and kamacite, and of their eutectoid (?) aggregate, plessite. By analogy, one would infer the existence of similar phases in the iron-manganese alloy just described, such an inference being strengthened by the similarity displayed in many ways by nickel and manganese in their effects on iron. Iron-manganese Alloys are Regarded as Homogeneous Heterogeneity of iron-manganese alloys has not been recognized heretofore, judging from the data of standard metallographic books and those of original papers. In fact, iron and manganese are quite generally regarded as completely miscible in the solid state. Dr. Rosenhain,2 for example, writes: Alloys of group (a), of which those of gold-silver, iron-manganese, and copper-nickel are examples, when allowed to crystallize sufficiently slowly to attain their condition of final equilibrium, present a microstructure exactly similar to that of a pure metal—the homogeneous character of the molten solution is in that case so completely maintained that the microscope cannot detect the presence of any second constituent. Guillet and Portevin3 refer to the iron-manganese system as an example of the "case of metals completely miscible in the solid state."

By Walter Mathesius

The use of high blast heats on furnaces melting Mesaba ores is still the exception, the average blast temperatures carried on Mesaba stacks seldom reaching 1,100" F. Some 15 years ago, when the use of fine Mesaba ores in larger proportions to the total burden first came into practice, it was found difficult to employ even those blast heats which it had been possible ' to carry with the Old Range ores. The furnaces refused to "take" high heats, and would persistently labor with irregular stock movement, which impaired the practice and increased the coke consumption. By lowering the blast temperature, the working of the furnace became smoother, and experience taught that with the softer and more easily reduced Mesaba ores, satisfactory practice could be obtained with comparatively low blast heats. Therefore a large number of Mesaba furnaces have ever since been operated on very low blast temperatures, ranging between 800" and 1,000" F. In the face of these difficulties, furnace men were slow to recognize that furnaces could be so designed as to permit a free movement of the charges in combination with the use of higher heats. Since higher blast temperatures did not seem to offer great advantages in the smelting of Mesaba ores, investments in hot-blast stoves were not as attractive to the furnace man as labor-saving devices and means of obtaining greater tonnages. The effect of this preference is keenly felt at the present time. It has brought about the peculiar situation, that the majority of our furnace plants (among them some of the most progressive) lack modern and efficient stove equipment. The result is that, even where the proper furnace lines have been adopted and conditions are now favorable to the successful employment of higher blast temperatures, the heats required for the greatest fuel economy are not available. In an endeavor to obtain results, even with inadequate equipment, we find that, at many plants, the stoves are worked far beyond their capacity. This leads, in many instances, to a situation in which the coke saving in the

Jan 1, 1915

By Fred E. Beasley

Metallurgists have, in the last ten years, overcome many difficulties of high zinc in lead blast-furnace slags. This problem was brought to the front at the close of the war, by the price obtained for lead and the high cost of labor, fuel, and flux. The Port Pirie plant of the Broken Hill Associated Smelters, Australia, early in 1919 had increased the zinc oxide in its blast-furnace slags from 13.5 to 20 per cent., at the same time reducing the iron oxide in the slag from 33.5 to 25.6 per cent. These results were obtained by the addition of 10 to 12 per cent. granulated slag to the roaster charge, improved conditions of the crushing, mixing, and roasting of the constituents of the sinter. By experiments and improvements, the practice at this plant has developed a normal slag that carries 22 per cent. zinc oxide, and at times runs as high as 33 per cent. zinc oxide, with only slight furnace trouble. The granulated slag is used not only as a diluent to the sulfur in the roaster charge, but on account of its value as a slag-forming material and the physical properties it imparts to the finished sinter. The ore is mixed with flux and given a double roast. All the first roasting is by sintering with the Dwight & Lloyd machines. The final roasting is performed partly by Dwight & Lloyd machines and partly by H. & H. pots. The charge to the blast furnace consists chiefly of the final sinter, slag shells, siliceous and non-bearing sulfur lead ores. High blast and ore columns are the practice. The greatest care is taken to obtain in the final sinter as low a sulfur content as possible. Attracted by the success of the Port Pirie plant, The Consolidated Mining & Smelting Co. of Canada, at Trail, B. C., in 1920 treated its crushed primary sinter with about 15 per cent. granulated slag. The Trail smelter at this time was short of ore and a large part of the charge to the roasters consisted of the zinc-plant residue having the following analysis: 12 to 16 per cent. lead, 12 to 16 per cent. zinc, and 25 to 30 per cent. iron. The lead in the final sinter was low and the physical condition was so bad that all the sinter had to be screened before it could be charged to the blast furnace. Also it disintegrated on being trammed and dumped

Jan 1, 1925

By B. F. Fackenthal

For the past year we have had in use at the Durham furnace a set of hollow pig-patterns made of iron, which have given such satis factory results that I think a description of them would be of interest to the members of the Institute. Before the introduction of these iron patterns, we used the ordi nary wooden patterns, made in the usual way with straps of iron along the sides to protect the corners from wearing off. The size and shape of these wooden patterns are shown in Fig. 1. When the patterns were new, the pigs weighed about 130 pounds each, and were uniform in size, and the moulding was clean. But constant use soon wore the patterns away; and at the end of a few months, the pigs had become, in consequence, so much lighter that the moulding, daily, of several additional beds was necessary to accommodate the same quantity of iron. As the patterns wear away, the moulding becomes ragged and the pigs rough. Moreover, wooden patterns are often destroyed by fire. I have known an entire set to be destroyed by a break-out of iron. The iron pattern which I have designed is free from all the objections to the wooden pattern, being not only more durable, but, in the long run, far cheaper. The iron pattern having the same shape and size as the wooden one, is made of the best flange-iron, of No. 13 gauge. After the sheets have been cut to the proper size, three heats are required for the flanging. At the first heat, each piece is stamped in a cast-iron form, which gives the proper shape to the bottom part of the pat-

Jan 1, 1889

By S. E. Bretherton

Mr. AllaW Gibb, of Mount Perry, Queensland, Australia, in an interesting and instructive paper,* describes fully the great difficulties metallurgists encounter in seeking to produce marketable copper by smelting refractory ores containing arsenic, antimony and bismuth. After reading this description, I prepared a paper for the Institute, which was, however, through a misunderstanding on my part, published in the San Francisco Mining and S'cie~ztiJic Press of April 4, 1903, after its acceptance by the Council, and without credit to the Institute. Consequently, it cannot be reprinted for the Transactions of the Institute; and I can only here summarize its contents and conclusions, referring the reader to the article above named for further particulars. In that article I omitted the col~sideration of bismuth, which we do not encounter in our smelting at Val Verde, and added lead and zinc, which Mr. Gibb had not discussed; and I undertook to shorn, by the record of a year's practice of the Val Verde Copper Co., Arizona, that copper-ores containing arsenic, antimony, lead and zinc may be successfully smelted, without preliminary roasting, in an ordinary blast-furnace with hotblast, making a normal slag, such as one would aim to make under ordinary conditions with cold-blast, so as to produce a marketable matte, containing not less than 45 per cent. of copper, free from arsenic and antimony, and below the limit in lead and zinc penalized by the refineries.

Jan 1, 1904

By S. LeFevre

, The solution of the housing and sanitation problem in mining communities, keeping in view both economic and humanitarian aspects, demands the best thought of the management of such enterprises. Upon the solution depend the health and comfort of the employee and his family, and as a consequence his contentment and efficiency. The following description covcrs briefly the methods installed for Witherbee, Sherman & Co. The operation of the system has been for about two years under the direct supervision of T. A. Hammond, Land Agent, and I am indebted to him for a compilation of the results. Sanitation The sanitary service covers 238 tenements. The installation cost was about $22 per tenement, as follows: Privy, $15; privy box, $3; proportion of incinerator cost, $2; garbage can, $2. The operating cost is $1 per month for each tenement. In considering the installation of an ordinary village system of water supply and sewerage two difficulties were encountered: First, lack of a sufficient quantity of water; there is no stream large enough, as Mineville is situated very close to the divide between the Lake Champlain and Hudson River watersheds. A storage reservoir was tried, but the only location available, among glacial hills, would not hold water. Second, the expense of digging ditches for the distribution of water and sewer pipes is so great as to be prohibitive; Mineville is built chiefly in a valley covered in the bottom with drift, in this case a combination of boulders and hard pan. Digging in this material costs 80c. to $1 per cubic yard, and pipes must be placed 41/2 ft. down to be below frost. The houses are scattered along 4 miles of streets, the oldest having been built in the original forest 65 years ago. The newer houses have gone up a few at a time wherevcr a clear space could be found and the slopes were not too steep. This has advantages, as it separates the dwellings into

Jan 1, 1915

By Oliver C. Ralston

A definite field of usefulness has developed for the brine-leaching processes of removing lead from ores and other products, so this paper reviews the developments, both in practice and in research, that have taken place since the publication of two earlier papers on this subject.' Several commercial plants are now in operation and a number of semi-commercial test plants have been operated in an effort to determine the limitations of brine leaching and to compare the results with other forms of treatment. Briefly, the brine leaching of lead depends on the fact that lead chloride and lead sulfate are soluble in saturated sodium-chloride solutions. In dilute brines, the solubility of the lead salt is less than in pure water; but in the strong brines, the solubility is greater than in pure water due, supposedly, to the formation of a double sodium-lead chloride. This is seen in Fig. 1, which shows the results of M. Demassieux2 on the solubility of lead chloride in sodium-chloride solutions of different concentrations and at three temperatures. The curves are not those given by Demassieux but are calculated from his results in order to convert them to units common in metallurgical work. The right-hand branch of each curve shows that if the solution is kept in contact with solid sodium chloride the lead chloride can be almost entirely displaced from solution, but that brines which are saturated with sodium chloride, but are not in contact with the solid salt, have a very high solubility for lead chloride. The diagram gives the analysis of the solution in grams of lead per liter, although the substance actually in solution is lead chloride. At 100" C. as much as 100 gm. of lead per liter can be obtained; but at lower tem-

Jan 1, 1924

By Robert P. Burrows

In preparing this paper the object has been to set forth facts relating to illumination problems, which, judging from the results realized in the iron and steel and other industries somewhat similar to mining, will tend toward furthering safety, production, and contentment of employees, as well as economy of operation in mines. By applying the principles of illumination with the assistance of modern appliances, the full benefits in efficiency may be derived from improvements already made in other details of mine operation. The lighting of a typical coal mine may be*divided into four distinct parts: 1, The lighting of the buildings about the top; 2, the lighting of the working faces; 3, general illumination at the bottom; and 4, special applications of lighting. The lighting of buildings about the top may be treated in the same manner as that of any other industrial plant, for we have a boiler room, an engine and generator room, a forge, a machine shop, and a hoist room. These can be well and efficiently lighted by the use of 100-watt tungsten-filament multiple lamps with proper reflectors so spaced and suspended that a power consumption of from ¼ watt per square foot in the boiler room to 1 watt per square foot in the machine shop is obtained. The methods that apply to this kind of lighting have been ably treated by a number of authors,' and for this reason a detailed discussion is unnecessary. The lighting of the working faces is usually done by means of portable lamps, of which there are four types in use: The oil torch, the acetylene Iamp, and the oil and the electric safety lamps. The different types have been fully described in numerous papers and articles and will not be covered here, although a few figures on the cost of operation will no doubt be

Jan 1, 1916

By Owen R. Rice

Changes in coke hardness affect the working of the blast furnace, for soft coke is an obstacle to proper furnace operation. Soft coke is due to a low hydrogen-oxygen ratio in the coal charged; increasing this ratio increases the hardness of the coke. The best means of determining the hardness of coke is the combined tumbler and ball-mill. What is Good Coke? It is not always appreciated that the correct answer may be such a reliable forecast to the physical condition of the blast furnace, and incidentally its entire operation, that among the ordinary variations in other raw materials and also in operating conditions, it assumes a position of primary significance. Inasmuch as coke forms two-thirds of the total bulk of material charged, this assertion will not seem exaggerated, and in the subsequent discussion its applicability will be presented. Coke is the ideal blast-furnace fuel because it is strong and hard enough to resist the crushing and grinding effects of the furnace burden, and its porosity allows the ready passage of gas upward through the charge, and rapid combustion before the tuyeres. Coke deprived of its strength and with a cell structure physically frail and chemically susceptible to solution by carbon dioxide has its chief asset eliminated. When the melting zones are approached the fluid cinder trickling toward the hearth will envelope the small pieces of coke and form a sticky mass that will retard the free passage of gas. Blast pressure will then rise, driving will slow up and fragments of coke will lie inert before the tuyeres, rendered practically incombustible by their slaggy envelop. It is believed also that coke messes at cast are attributable to quantities of such coke floating unburned in the hearth. It has been correctly claimed that coke rendered small outside the oven possesses greater combustibility than that so developed in the

Jan 1, 1922

By Owen R. Rice

Changes in coke hardness affect the working of the blast furnace, for soft coke is an obstacle to proper furnace operation. Soft coke is due to a low hydrogen-oxygen ratio in the coal charged; increasing this ratio increases the hardness of the coke. The best means of determining the hardness of coke is the combined tumbler and ball-mill. What is Good Coke? It is not always appreciated that the correct answer may be such a reliable forecast to the physical condition of the blast furnace, and incidentally its entire operation, that among the ordinary variations in other raw materials and also in operating conditions, it assumes a position of primary significance. Inasmuch as coke forms two-thirds of the total bulk of material charged, this assertion will not seem exaggerated, and in the subsequent discussion its applicability will be presented. Coke is the ideal blast-furnace fuel because it is strong and hard enough to resist the crushing and grinding effects of the furnace burden, and its porosity allows the ready passage of gas upward through the charge, and rapid combustion before the tuyeres. Coke deprived of its strength and with a cell structure physically frail and chemically susceptible to solution by carbon dioxide has its chief asset eliminated. When the melting zones are approached the fluid cinder trickling toward the hearth will envelope the small pieces of coke and form a sticky mass that will retard the free passage of gas. Blast pressure will then rise, driving will slow up and fragments of coke will lie inert before the tuyeres, rendered practically incombustible by their slaggy envelop. It is believed also that coke messes at cast are attributable to quantities of such coke floating unburned in the hearth. It has been correctly claimed that coke rendered small outside the oven possesses greater combustibility than that so developed in the

Jan 1, 1922

By Raymond B. Ladoo

TIIe milling of talc, as is the case with many non-metallic minerals, until recently, has not received adequate technical consideration, for the talc industry has become of importance only within the last decade. At first, talc was used only in the massive form, for foot warmers, griddles, and so on, and milling methods were unnecessary. As the demand for ground talc increased, producers adopted the machinery used in the milling of flour; in fact, many talc mills were rebuilt flour mills. Improvement has been slow, but today several types of grinding and separating machinery are in use; many of them, however, are still inadequate and inefficient. To determine the best methods of milling talc, it is necessary to understand the essential properties a talc must possess to fit it for a particular use. For toilet purposes, whiteness, freedom from lime and grit, fineness of grain, and good "slip" are essential. As a paper filler or coating, the talc must be white, uniform, and fine grained, have a good slip, and be free from grit and iron; freedom from lime is a disputed point. Fibrous talc is supposed to be superior to massive on account of the interlocking of grains in the paper, thus increasing its strength; this point, though, is in doubt. In general, talc for the paper industry need not be of as high quality or as fine grained as that used for toilet purposes. Talc is usually bought by sample, for which reason it is difficult to state the essential properties for use in paint, rubber, roofing, and so on. The adoption of standard screen tests and standard grades by producers is necessary before an accurate basis for manufacturing or selling standards can be established. Off-color talcs might be utilized, as in Germany and Austria, by the establishment of standard grades of colored talcs. The machinery to be used depends on whether the talc is fibrous, foliated, or massive; hard or soft; of uniform or variable grade; pure or impure. In some cases, it is possible to change the mining practice or to utilize different sections of the deposit in order to vary some of these factors, but in many cases these factors are fixed, so the milling must be designed to suit the conditions.

Jan 1, 1922

By P. H. Dudley

When we see the magnificent passenger-trains of from 8 to 12 coaches, drawn by locomotives weighing from 100 to 110 tons, at speeds of from 50 to 60 miles per hour between terminals, to make a schedule of 45 miles per hour, and freighttrains of from 50 to 75 cars of 60,000 pounds capacity, drawn by one locomotive, it is hard to realize that it lacks a few months of fifteen years since my pioneer 5-inch 80-pound rail for the United States was laid by the New York Central aid Hudson River Railroad Co., in July, 1884, on the Harlem line. Shorter passenger-trains, of from 4 to 5 coaches, are run much faster, notably the Empire State Express, the schedule of which for the 440 miles between New York and Buffalo requires 53.3 miles per hour, calling for a running-speed, for much of the time, of from 65 to 75 miles per hour, while higher speeds are very common. On other railroads laid with my 5 1/8-inch 80-pound rail, speeds of from 60 to 75 miles per hour obtain in daily practice. On many of the important railroads east of the Mississippi high speeds of from 60 to 70 miles per hour are part of the daily service. So many instances of speeds of 90, and even 100, miles per hour have been recorded that they must be considered as likely to occur on important lines, and provision must be made for them in the track. Solid mail-trains of from 6 to 8 cars are run from New York to Chicago, 1000 miles, in twenty-four hours; and, commeneing with the present year, shorter trains are running from Chicago to San Francisco, 2000 miles, in three and one-half days. In the freight-service equally important progress has been

Jan 1, 1900

By Eckley B. Coxe

In making surveys in the anthracite coal regions of Pennsylvania, the ordinary engineer's chain (50 or 100 feet long) is generally used, both above and below ground. Sometimes, where it is difficult to chain, as, for instance, across a chasm, a wire is stretched from one station to the other, the distance is marked on the wire and its length is then measured with the ordinary chain. Having had occasion lately to make some surveys where it was necessary to determine with great accuracy the position of the land or property line, not

By M. E. Lombardi

ThE Coalinga oil field is located on the west side of the San Joaquin Valley, California. The structure is in general a monocline, the edges of the oil horizon resting on the foot hills and dipping gently toward the east. One prominent anticline occurs plunging southeast. The earlier drilling was done in the foot hills comparatively near the outcrop and the wells were shallow. The sands were followed eastward and, in the case of the anticline, along the plunge, the wells becoming deeper and deeper until the depth of 4,000 ft. was reached and passed. There is nothing to show that the oil will not be found in quantity at still greater depth. In fact, some of the best producers have tapped the sand at close to 4,000 ft. The recovery of oil still farther to the east, and therefore at greater depth, seems to be mainly a question of drilling. In this territory the formations drilled through are chiefly sands and shales; they will not "stand up" in an open drilling hole; the casing has to be carried close to the bit, and it is always difficult to keep the casing free for any considerable distance. Ability to carry casing of comparatively large diameter without conductor pipes for distances of 2,000 or 3,000 ft. or over is desirable in such territory chiefly for two reasons. It makes it possible to enter the oil sand with a pipe of ample diameter; it eliminates one or more expensive strings of casing which act only as conductors for the water string, and furthermore, in territory where waters are encountered which corrode steel rapidly, it makes possible the construction of a rust- and alkali-resisting water string. It is always desirable to shut off top waters, which may lie within 100 ft. or less of the oil sand, with 10-in. pipe. Where the depth is so great that a practical weight of 10-in. pipe will not withstand the probable collapsing pressures, 81/4 in. at least is desirable. About the limit of rotary drilling to date in California seems to be the setting of the 10-in. string at 3,200 ft., although the rapid advance in rotary work during the past year seems to indicate that this depth may soon be increased. It is my purpose now, however, to treat only of cable-tool drilling. The problem is to reach a depth of 4,000 ft. or more with a string of

Jan 1, 1915

By J. P. Dovel

Having been requested to prepare a paper referring especially to my patents as applied to blast furnaces, I shall confine my discussion to those improvements and inventions pertaining directly to the blast furnace itself, omitting a number of patents pertaining to blast furnace operations. I spent the early part of my life in the construction of blast furnaces, from rivet-heating to drafting and designing, much of the work relating to repairs and rebuilding. Later, as I became more interested in maintenance and operation of a group of blast furnaces, one fact especially made a lasting impression on me; namely, that a very expensive piece of equipment was very short lived, and the expense of replacement, relining, etc., in many instances consumed most of the profits made while in actual operation. As far back as 1910, I began to lay plans to build with a view to permanency and not merely for a blast, the common expression at that time, which really meant only one or two years in southern practice. Since we did not then blow so hard most failures occurred in the bosh and hearth. The failure of either generally meant a complete new lining, as the inwall did not look good enough to last another blast. At that time we were using two types of hearth, one of heavy steel plate protected by well and sprays on the outside, and one of heavy cast iron, made in sections and bound together with wrought-iron bands. The former type proved dangerous, for if the iron should come through, as it frequently did, it would cut away the steel plate, quickly turning large quantities of molten iron into the water, causing a severe explosion and costly repairs. The cast-iron type gave trouble also by the iron finding its way through the joints between the sections, which was not so serious, since the iron would not cut a way the heavy cast-iron sections very quickly, and generally caused only small damage, and repairs were easily made. Both types have been improved; the former by the introduction of cast-iron water-cooled plates, placed on the inside of the steel jacket. This has proved to be very good construction, but has no real claim to permanency. As the cast-iron plates cannot be replaced from the outside, the operator must assume that some time, at no late date, he must replace them from the inside, which will mean a blowing out of the furnace. The heavy cast-iron hearth has been improved by making the