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By Carl G. Hogberg

During recent months, the acute shortage of steel scrap has necessitated the use of higher percentages of hot metal in the open-hearth charge. With these higher percentages, the sulphur content of hot metal has become an increasingly important factor in the control of the quality of steel. Thus, the blast-furnace operator is faced with a demand not only for more output but for a higher quality of product. More than go per cent of the sulphur charged into the blast furnace is contained in the coke. With the tremendously increased demand, the trend of coke quality has been toward higher sulphur, higher and more variable ash, and poorer physical characteristics. This has been brought about by less selective coal mining, increased production rates in coke plants and resumption of operations of many beehive ovens. In a single month last year, one blast-furnace plant in the Pittsburgh district operated with coke obtained from 52 sources. The effect of variability and quantity of coke ash on blast-furnace operation is well known. The improved results obtained with furnaces operating on coke made from washed coals compared with results on coke made from raw coals is ample proof of the effectiveness of low and uniform ash. Coal-washing plants, however, require a large capital expenditure, and are impractical for most of the large number of small producers, whose operations are confined chiefly to periods of abnormally high demand. Furthermore, not all coals are physically adaptable to washing. Principles 01 Iron Desulphurization in Furnace Although investigations of the blastfurnace process have not clearly revealed the detailed mechanism of iron desulphuri-zation within the furnace, they have been sufficiently extensive to permit a postula tion of the following general principles: 1. Iron, in its travel from furnace top to hearth, continues to pick up sulphur until it has reached tuyere level. The zone of iron desulphurizatioil is relatively narrow, being confined to a vertical distance extending zapproximately from tuyere level to the iron-slag interface in the hearth.' 2. Combustion of coke is confined to spherical zones extending approximately 40 in. from the nose of each tuyere.= Coke ash is released as the coke bums in these combustion areas. 3. With a constant amount of sulphur in the charge, the sulphur content of the metal product is a function of hearth temperature, final slag analysis, slag viscosity, and slag volume. Both hearth temperature and slag analysis affect slag viscosity. For a slag of constant analysis, viscosity always decreases with an increase in temperature;

Jan 1, 1943

By William Colby

INCREASING governmental control of human activities seems inevitable. Within the present generation railroad rates and the public sale of water, gas and electricity have been subjected to rigid regulation. The field of control is slowly but surely being enlarged, and it is the object of this paper to outline in brief space this tendency toward increased governmental regulation, as far as mineral resources are concerned. Our, dual form of government complicates the problem. We must deal not only with the powers of the Federal Government, but also with those of the individual States. In order to approach this question intelligently we must have some conception of the powers and jurisdiction of the States and the United States, respectively, over mineral lands and mineral production. Starting with the Federal Government we find it never owned any public domain within the States that formed a part of the thirteen original colonies and, for historical reasons, the same is true of Tennessee and Texas. The Mid-Western States were erected out of the public domain which was acquired by the United States through cession from the original States or by purchase from foreign nations. Congress early in the last century adopted a policy of leasing public lands containing lead deposits in a portion of this Mid-Western territory, but later abandoned that policy and disposed of such public, lands, mineral and non-mineral, outright so that they long since have practically all passed into private ownership and become subject to the general jurisdiction of the respective State governments. In the great public domain lying west of the Mississippi River, and in portions of some of the Southern States, the Federal Government adopted a new mineral land policy following the discovery of gold in California. (Acts of Congress of 1866, 1870, and 1872.) The mineral lands within these States, instead of being disposed of on the game basis as agricultural lands, as occurred in the Mid-Western States, were segregated from agricultural lands and sold under special laws intended to encourage mining, giving the miner the free right to locate and work these mineral lands without paying any royalty to the Federal Government. A miner could also obtain at slight expense a complete

Jan 1, 1930

By Fred C. Bond

AN improved method of measuring the surface area of a comminution product down to any desired particle size has been developed. The method is largely graphical, and requires relatively little calculation. It can be used with materials containing large amounts of clay, or excess fines of any kind. The surface area S is expressed in terms of square meters per100 c.c. of solids. This expression is based upon volume rather than weight, and is considered to be more suitable for crushing and grinding problems, where the machines ordinarily are operated under constant volume conditions. It has the advantage that surface areas of different materials can be compared directly, even though their specific gravities may be unknown. The "specific surface" of a material, or the number of square centimeters of surface area per gram, is equal to 100S divided by the specific gravity. Particle diameters are expressed in microns, K being the diameter of the largest particle present and L the grind limit, or lower size limit of the particles formed directly by grinding. The value of K does not always coincide exactly with the diameters of the largest particles retained on the coarsest testing sieve, since the largest particles formed in crushing and grinding are often slabby, or have other irregular shapes. L has been chosen as 0.700 microns, or number 27.5 in the size number series,1 for reasons to be explained later. S is the total surface area of all particles larger than L microns in diameter. When a homogeneous comminution product is separated into fractions whose diameters bear a constant geometrical relationship [p] to each other, and the log per cent weight retained upon each size and passing the next larger size is plotted against the log diameter, a straight line results, which is called the distribution line, and which has a slope in called the distribution modulus. The distribution line can be extended to any desired limiting size, and the total surface area of all particles larger than that size can be calculated. The equation of the distribution line is w = Cxm [I] where w is the per cent weight retained on each successive sieve; x is the particle diameter; and C is the per cent weight retained on an opening of unit diameter.2 Theoretically, the distribution line can be extended to include infinitely small particles. When it is so extended, and the assumption is made that no fine particles are present other than those accounted for by the equation; the log per cent cumulative weight passing any series of sizes, plotted against the log diameter passed, will result in a straight line lying parallel to and above the plotted distribution line.

Jan 1, 1941

By Haakon Styri

In discussion of Dr. Fitterer's paper on electrolytic separation of slag inclusions, some results from experiments on electrolyzing high-carbon steels at the SKF Research Laboratory were given. l We had been working for a number of years on chemical methods of separating the inclusions but could not find any solvent that would dissolve the steel without attacking some of the inclusions, particularly the sulfides in which we were mostly interested at the time, nor any that would dissolve all sulfide or silica in filings as fine as 300 mesh without attacking the iron. Electrolytic methods had also been tried without any real progress until we learned from the Bureau of Mines about the use of the collodion membrane around the anode. Our early difficulties were to some extent different from those of the Bureau of Mines, perhaps because of experimenting on steels of higher carbon content, but most of our difficulties were of the same nature as those of the Bureau. Formation of large quantities of basic hydroxides was traced to the use of impure and oxidized ferrous sulfate. To protect from oxidation from the atmosphere during electrolysis, after purer chemicals were obtained, a thin layer, 54 in. or less, of molten paraffin was poured over the surface of the cathode liquid, where it solidified. It was found desirable also to hold the cathode chamber slightly acid by initial addition of 0.1 per cent acid (sulfuric or hydrochloric) to prevent precipitation of the hydroxide. This acidity would decrease gradually and the cathode chamber would become neutral before the end of the electrolyzing. The increase in acidity in the anode chamber during our early experiments had appeared to cause attack of the precipitated sulfides in the residue, which originally was analyzed by volumetric method for sulfur, and for this reason a 10 per cent sodium citrate solution was substituted for the sulfate in the anode chamber. With the use of the more reliable Eschka method for sulfur determination (as in coke), there was however no serious attack on the sulfides in the anode chamber with a variety of different liquids, as shown in Tables 2 and 4. The Eschka method gives a higher sulfur value than

Jan 1, 1933

By David P. Hale

THE Cartersville barite district is near Cartersville, Ga., in the southeastern part of Bartow County, about 43 miles northeast of Atlanta. The area over which active mining is being done extends about six miles in a northeasterly direction and is about three miles wide. Barite production in Georgia began about 1907,1but the existence of the mineral was known much earlier. At present four companies are operating washing plants and three other plants are operating on leases from those companies. The chief impurities in the ores are silica and iron oxide. The silica is easily washed and jigged out and gives little trouble. Any gravity separation between the iron and the barite is generally poor, owing to the slight difference in specific gravity. Two operators have turned to magnetic separation as the solution to the problem and have made remarkable improvements in the finished product. By very close sizing prior to jigging, the separation between iron and barite has been greatly improved. In some localities manganese occasionally appears as an impurity, but is fairly easy to eliminate by gravity methods. A considerable percentage of iron and silica is eliminated by being broken free in crushing. Until methods for the recovery of fine ore were introduced a minimum of crushing was done, but modern methods allow more than formerly. The depression years brought about the need for increased recovery of ore handled on the part of the producers and a demand for a higher grade of product on the part of the buyers. Improvements in milling and mining practice have had a far greater effect, however, than merely the production of a better grade of ore with greater recovery. They have opened for mining large areas that a few years ago were considered noncommercial deposits because of their inaccessibility and their high iron content. It is no exaggeration to say that the new methods have more than doubled present known reserves of barite ore of economic value.

Jan 1, 1938

By Charles T. Holland

In connection with the problems of bumps in coal mines, much has been written concerning the manner in which roof action and methods of mining enter into the pressure effects observed but little has been written upon the relationship between the physical properties of coal and associated rocks and the cause of bumps. The question of the strength of bituminous coal, of its elasticity, and the energy it can absorb and store has been neglected to a considerable extent in connection with this problem. It is generally agreed that one of the necessary conditions for bumps is a strong coal.' Therefore, since the strength of coal is an important factor, a study of this property should be made. several investigators have conducted laboratory experiments upon the strength of coal, using small cubical or rectangular test pieces. An attempt will not be made to go into all of these investigations, but they are briefly summarized in Table I. Experiments of the U. S. Bureau of Mines and of Lawall and Holland (pp. 27 and 34 of ref. 9), however, indicate that regardless of whether the coal is a soft and weak semihituminous (Pocahontas No 3) or a hard and strong bituminous (Stockton) coal, the maximum ultimate unit stress developed in laboratory tests is higher in small cubes than in larger cubes of the Same coal.

Jan 1, 1942

By J. H. Nead

OBJECT THE experiments herein described were undertaken with a view to investigating thoroughly the influence of carbon on the tensile and impact physical properties of carbon steel. The original comprehensive plan included investigation of a series of steels with varying carbon contents over a wide range of heat treatments. These were to be studied as to their static tensile qualities, and also their resistance to shock as measured by the Charpy impact bending test. A thorough metallographic examination was likewise included in the schedule. The Charpy tests and the metallographic examination have not been completed to date, and are accordingly withheld for another report. All heat treatments, however, have been made, and the static tensile tests have been completed. A description of the procedure of the experiment and the results thus far obtained are therefore reported herein. CONCLUSIONS No attempt is made to draw general conclusions from these experiments, but the curves and the data are submitted as being of general importance and wide applicability.

Jan 12, 1915

By Frederick Prime

The Great or Cumberland Valley, which (under a variety of names) extends from Canada, through Vermont, Massachusetts, Connecticut, New York, Pennsylvania, Maryland, Virginia, and East Tennessee, to Alabama, is composed in part of limestones lying immediately above the Potsdam sandstone. Geologically, they belong to the era between the Potsdam and Trenton, and are commonly termed Calciferous. A peculiarity of these limestones is their great richness in limonite or brown hematite ores, which form a large part of the charges of the many blast-furnaces situated on their outcrops. The mines from which the ores are obtained are (properly speaking) only pits or quarries, from which 'the ore is extracted by means of picks or shovels, powder being rarely employed. The ore does not occur casually at any point in the limestone, but forms regular lines, following apparently the outcrops of certain beds. In Lehigh County, Penn., where 1 have been engaged on the Geological Survey of the State, during the past season, there are three or four of these lines. During the progress of the survey I was much struck by the fact that the two most important and promising lines of outcrop were, one at the base of the crystalline schists (forming the South or Lehigh Mountain), and overlying the Potsdam sandstone conformably; the other, near the line of junction of the calciferous limestone with the slates of the Cincinnati epoch. The mines along these two lines, following the topography of the country, were in place, richer and more permanent than those of the belts between, which were generally leaner. A peculiarity of these two lines of outcrop, and one to which I particularly desire to draw attention, is the occurrence of what I at first mistook for a highly altered slaty limestone, but which subsequent analysis proved to consist in great part of damourite (hydrated potash mica). The following are analyses of this damourite slate: I. From Fogelsville, Lehigh County, within a few hundred feet of the contact of limestone and slate, by Dr. F. A. Genth. 11. From Hensingerville, Lehigh County, within 300 feet of gneiss, by Mr. Sydney Castle, of the University of Pennsylvania. 111. From Allentown, Lehigh County, within 150 feet of con

By H. Kreye, E. Hornbogen

An increase in electrical resistivity during aging or after quenching from high temperatures is known in a large number of alloys. Plastic deformation of alloys in this condition leads to a decrease in resistivity. This anomaly has been called the K effect. This effect has been reported in Ni-A1 alloys by several authors.'-4 The nickel-rich solid solution y is in equilibrium with Ni,A1 (y') with CkAu structure. With this alloy it is therefore possible to produce a homogeneous solid solution as well as a coherent mixture of y and y. The changes in resistivity should be determined for these two conditions. In the present literature on the K effect no distinction is made between them, because of experimental difficulties to determine small coherent particles. In the present work transmission electron microscopy and image formation by (100) and (110) superlattice reflection of the y' phase was employed for a detailed determination of the temperature dependence of solubility of aluminum in nickel and the resulting microstructures with alloys up to 15 at. pct Al. The solubility line follows the equation indicating a smaller solubility than given in the literature at temperatures below 700°C. X-ray scattering was used to get evidence for short-range ordering in the homogeneous solid solutions.5 Parallel to the determination of the microstructure electrical resistivity was measured as a function of aging tempera- ture, time, and aluminum content, which led to the following results: 1) Resistivity increases slightly during aging of a homogeneous solid solution which was quenched from a high temperature as a consequence of short-range order, Fig. 1, curve c,. 2) Resistivity shows a maximum during aging of all alloys which are heterogeneous according to the phase diagram, but which were obtained as a supersaturated solid solution by quenching, Fig. 1, curve c,. The increase in resistivity is mainly due to coherent precipitation of y'. In Fig. 2 the effects due to short-range order pN and due to precipitation of particles APT are shown for an aging temperature of 450°C and different aluminum contents. 3) The increase in resistivity depends on the size !f the coherent particles. A critical diameter of 15 5A for the maximum was found for all alloys and aging treatments by extrapolating particle sizes determined by electron microscopy. 4) The maximum of resistivity of the heterogeneous conditions does not occur if particles, larger than the critical size, have formed already during quenching. 5) Plastic deformation destroys short-range order as well as small coherent particles. Therefore a decrease in resistivity is observed if the increase in resistivity caused originally by these phenomena is

Jan 1, 1967

By F. M. Lewis, J. F. Meyers

This second progress report of grinding presents comments regarding ball consumption and data pertaining to the hydroscillator, which is closed circuited with the tricone mill. A study and postulate of how balls function is presented. The paper summarizes cost reduction obtained by the method of "grinding for the tail race" rather than for "mesh tons." AT the Regional meeting in Columbus, Ohio, in September 1949, the authors presented a progress report of the first year's operation with a Hardinge tricone mill in closed circuit with a Dorr hydroscillator. The present report covers our findings on this grinding circuit to January, 1950. In order to clarify our position, the authors wish to state that no invention or discovery is claimed. The Tennessee grinding circuit is simply an engineered arrangement of known equipment arranged to incorporate all of the recognized efficiency factors of comminution as they are known today. As we concluded our first report, it was not clear how effective the Tennessee mode of operation would be on harder ores. The only information we had at that time was that the Bond grindability test at 48-mesh gave 6.04 g of undersize per revolution. A later test on the same sample at 100-mesh reduced this to 2.04 g, which is definitely in the harder ore class, and which makes the reported 8.80 kw-hr per ton of —200-mesh produced a very creditable figure. Hydroscillator Operation The machine as now developed has proved dependable and easy to operate. It starts easily under full load and responds quickly to operating adjustments. The hydraulic holes in the oscillating plate do not plug when shut down under full load. Absolutely clean hydraulic water is essential. An Elliott water screen or similar device should be an essential part of the machine to keep all scale and trash of all kinds from the hydraulic compartment. The shift operators have expressed approval of the machine. In table I we show a typical screen analysis of the hydroscillator rake sand as compared to conventional classifier rake sand on our ore. We have previously reported that of the 28.7 pct reduction in power, the hydroscillator contributed 6 pct, on a —200-mesh basis. This is a conservative figure. Some of our test runs indicated that the saving was as high as 10 pct. Laboratory studies by the Dorr Co. engineers indicate that still greater efficiency is possible. Nevertheless, the 10-ft plate is overflowing 2100 tons per day at 3 to 4 pct + 65-mesh, which is a very creditable performance by

Jan 1, 1951

By J. L. Jernegan

IN a former paper laid before the Institute, entitled Lead and Silver Smelting in Chicago, I endeavored to give a description of the manner in which argentiferous lead ores from the far West were treated in that city, and also described the method employed in the desilverization of the silver-lead produced from the ores. It will perhaps be remembered, that very siliceous ores, principally from the Emma mine, Utah, containing over 40 per cent. of silicic acid, were then being treated at the works of the Chicago Silver Smelting and Refining Company in the reverberatory furnace, with the use of metallic iron, in the form of borings, as a precipitating medium for the lead ; and that the extraction of the silver from the resulting bullion was effected by the Balbach process. The Swansea Works, also the works at Brighton, just outside of this city, and those of S. P. Lunt, were at that time treating ore of a similar character by the same method. The inadequacy of the process for the ores then under treatment was, in a manner, pointed out by me in the paper above referred to. Since the time of its writing, many and important improvements have been made in the metallurgical treatment of Western ores in Chicago. Most ores are now roasted previous to reduction, and the latter operation is conducted in the shaft-furnace instead of the reverberatory. Such is the fact at least at the Swansea Works, the largest in the city, and here it was where the first steps were taken in this direction. In selecting these works as the subject of this paper, I am in hopes that the description given of the various metallurgical operations, as they are there now conducted, will serve to give an idea of the improvements introduced

Jan 1, 1876

By M. E. Wadsworth

The chief value of a paper like this consists in its statistical tables, putting on record material useful to future inquirers. The data here given have been compiled from time to time as far back as 1874, and subsequently brought up to date. Much of the data could only be procured by personal inspection of the catalogues, as for example those relating to the Lawrence Scientific School, which were taken at that time from the catalogue of Harvard University, on file in the University Library, by the present writer, then an instructor in that institution. These tables of statistics were first prepared for publication by the present writer in 1888-1890 and published in his "Report to the Board of Control of the Michigan College of Mines " (Michigan Mining School) for 1886 to 1890. Later they were revised, enlarged and brought down to date in his reports to the same Board for 1890-92; and since that time they have been, at his request, revised, greatly enlarged and kept up to date by Mrs. Frances H. Scott, the efficient Librarian and Secretary of the Michigan College of Mines. Great credit is due to her for her devotion to this work and her untiring efforts to have the statistics complete as far as possible. The accompanying running commentary is given simply as a string to hold these statistics together, and in order to call attention to a few points. In compiling any tables like those accompanying this paper, one is confronted by serious difficulties, some of which render these results only approximate. In order to insure greater correctness and fullness in the statistics to be published the writer's reports for 1890-92 were sent to the president of every college, university, and technical school in the country, so far as known. Accompanying this was a personal letter which asked of each one his aid in correcting any errors, and in giving fuller statistics. But relatively few answers were obtained, and subsequently from one to many letters have been sent to each institution not formerly reply-

Jan 1, 1898

By W. H. Emmons

This paper is the third of a series treating of the relations of ores of the metals to igneous rocks. The first' outlined the general problem; the second2 proposed a classification of lode ores, and showed how the distribution of the deposits is in a certain measure controlled by the contour of the roof of the parent batholith. A number of mining districts were discussed, the examples being selected chiefly from group 2 of the classification. The proposed classification of lode deposits, briefly stated, is as follows: 1. Cryptobatholithic: near hidden batholiths which have not yet been exhumed by erosion. 2. Acrobatholithic: in and near cupolas or domes of batholiths—the high points of larger underlying masses. 3. Epibatholithic: on batholiths, near their rims, but below eroded cupolas, or below the highest points of the roofs. 4. Embatholithic: among closely-spaced outcrops which probably are parts of a single bathohth. The outcrops of invaded rocks predominate. 5. Endobatholithic. in and near roof pendants of large batholiths. The invading rocks predominate. 6. Hypobatholithic: in deeply-eroded batholiths, where even roof pendants have been eroded. Deposits of certain metals on the roofs of the batholiths are much more numerous in some positions than deposits of other metals, and the horizontal zonal arrangements of lodes of different metals are shown characteristically only in certain groups. In a large measure, also, the character of the fracturing and metallization is determined by the position on the roof of the batholith. The disseminated copper ores in porphyry constitute a clearly-defined type. All the deposits of these ores, as far as known, are in group 2. They appear to have been formed only in and around small upward projections of the parent granitic masses, and not low on the walls of the

Jan 1, 1927

By Peter F. Weiss

INTRODUCTION Austro-American Magnesite Co. of Radenthein, Carinthia, Austria, is a wholly owned subsidiary of General Refractories Company, Bala Cynwyd, USA, and has been incorporated since 1908. The raw material basis is the magnesite mine on Millstatt Alp, located in the mountains near Radenthein at an altitude of 1500 m above sea level. The mined crude magnesite is crushed in a surface crushing plant to minus 35 mm and consequently hauled to the processing plant in the valley by means of a cable conveyor of 7.5 km length. The annual production of 350,000 mt is accomplished with 20 supervisors and 122 workers (of those 60 are in the mine). The labor requirement for block caving is 0.2 man hr/mt, corresponding to an output of 40 mt/man/ shift. GEOLOGY The rock in this area mainly consists of crystalline type schist. There are mostly coarse crystalline garnet mica schists with varying mineral concentrations, and to a lesser degree, some deposits of amphibolite and marble. This petrographically rather homogeneous rock is divided into numerous different tectonic units (covers and scales). The magnesite deposit itself is a foreign body that also has been tectonically inserted. The ore body has a length of 600 m, is 450 m deep, and has a thickness of 60 m on average, varying substantially however, The contact with the accessory rock is very precise. The deposit 1 = ore body, 2 = previous open pit, 3 = caved in area, 4 = block under excavation, 5 = land slide consists of coarse crystalline magnesite in excess of 90 per cent. The few impurities consist of dolomite and talc. The lower part of the deposit is inclined parallel with the schists at approx. 40 degrees. In profile it can be seen, however, that the upper part of the deposit has been tilted over. This is a result of a post-glacial land slide reaching from the bottom of the valley at 1400 m above sea level to the peak of the mountain at 2100 m above sea level. Due to this land slide, with horizontal movements of up to 100 meters, as well as older tectonic events, the rock was crushed to a vast extent. Through mining, the land slide was reactivated; about 40 million cubic meters of rock slowly flow downhill almost like a plasticized mass. Near the surface, the rock is completely crushed and resembles a glacier rupture zone. This renders the originally impermeable rock completely water permeable so that practically all of the annual quantity of 1000 to 1100 mm of precipitation permeates the rock and flows down into the mine. The carbonate rock turns into the preferred aquifer. From 1700 to 2600 cubic meters of water have to be pumped from the mine daily. On one hand, additional problems are caused by the water in the mine, on the other hand we have learned to make use of the down- hill pressure forces to improve the crushing process during block caving. The boundary of the downhill shearing zone moves downward as mining goes deeper.

Jan 1, 1981

By C. F. Frey, P. J. Ahern, J. F. Wallace

THE high affinity of molten titanium for oxygen and nitrogen has resulted in considerable difficulty in developing a satisfactory melting procedure. It has been found necessary to perform melting operations either in a vacuum or inert atmos-phere, and failure to obtain a satisfactory crucible material has necessitated the use of special melting techniques. These techniques have included cold-mold arc-melting utilizing both consumable and nonconsumable electrodes, skull-melting, drip-melting, and levitation melting.' It has been recognized generally that the melting of both titanium sponge and titanium scrap would be expedited by the use of induction-melting procedures but the lack of a crucible capable of containing molten titanium without objectionable contamination thus far has prevented the use of induction-melting procedures. An extensive investigation: which included the common refractory materials was not successful in developing a satisfactory crucible for induction-melting. Some success in induction-melting has been attained with the use of a graphite crucible material, but in all instances sufficient carbon contamination was encountered to cause an appreciable reduction in the optimum mechanical properties of the titanium or titanium alloys. These investigations"" report an increase in the carbon content from approximately 0.3 to 1.0 pct. A serious reduction in the tensile and forming properties has been observed as a result of this carbon contamination. The development of a satisfactory induction-melting procedure for titanium without undue contamination undoubtedly would have considerable economic significance. Such a process would be particularly useful in melting the large quantities of low-cost scrap now available that can only be reprocessed at considerable expense and difficulty by the more elaborate techniques referred to previously. In addition, the process could provide a satisfactory method of melting and alloying of bri-quetted sponge for forging ingots.

Jan 1, 1959

By G. C. Eggleston, W. E. Keck, W. C. Lowry

The potential iron ores of Michigan can be classified from the stand-point of the predominant impurities into siliceous, sulphurous and phos-pllorous ores. Research on the flotation of each of these classes of ore was carried out in the following manner. The flotative properties of the valuable minerals of iron and of the worthless and deleterious ones were first determined under carefully controlled conditions. This information was then applied to synthetic mixtures of these minerals, which simulated the potential ores. Finally, data obtained during the two foregoing stages of the investigation were used in the flotation beneficiation of samples of the potential ores. Hematite is the most important mineral of iron in each of the above classes of iron ores. This paper deals with a study of that mineral in the very nearly pure state. There are two varieties of the mineral—specular hematite and massive hematite. The latter is the principal variety, but the former is an impor-tant' mineral of iron in some of the potential ores. Because of this condition, flotation research was performed with both varieties. This research consisted essentially of: (I) a survey of the literature on hematite flotation, (2) the preparation of pure massive hematite, (3) the preparation of pure specular hematite, (4) the development of a laboratory procedure and (5) flotation experimentation. For the purpose of this discussion, the results of previous work on the flotative properties of hematite can be summarized as follows: Specular hematite has been readily floated with terpineol and sodium oleate or heptylic acid1. Further, hematite has been likewise floated with a proprietary frother and oleic acid or other soluble oils2. Finally, lead nitrate, copper sulphate, silver nitrate, chromium nitrate, mercuric chloride, ferric

Jan 1, 1939

By R. C. Thompson

THE lead-zinc mill of Phelps Dodge Corporation, Copper Queen Branch, Mines Division, Bisbee, Arizona, is about 3 miles from the main hoisting shafts of the Junction and Campbell mines at Lowell. All the ore treated in the mill is obtained from these two mines. The mill was designed for an all-flotation treatment of 450 tons of lead-zinc ore per day, and has been operated at that capacity since the start of milling operations on Nov. 17, 1945. An expansion program is underway whereby the mill capacity will be doubled, but this article describes only the original program and presents results of the 45o-ton operation. GENERAL DESCRIPTION The milling plant consists of an ore-receiving bin and three steel-constructed buildings housing, respectively, the primary crushing equipment, the secondary crushing equipment, and the ore-storage bins, grinding, flotation, and filtering equipment. The mill site is the one that was used for a previous copper-concentrating plant, which was operated over a period of years and then dismantled. There remained from this operation some milling equipment, a reservoir for mill-water supply, railroad tracks, tailings-disposal ponds, and a steel-constructed building, all of which were incorporated in the design of the present lead-zinc mill. The location plan of the plant layout is shown in Fig I. The flowsheet of the mill is based on the conventional scheme for lead-zinc treatment, however, several innovations were made necessary in order to utilize the existing building. Placing of equipment also was governed by the design of the building; thus making the mill unique in many respects. The following brief description of the mill building as adapted to the flowsheet points out the compactness of the installation and some of its unusual features. The first floor (Fig 2) houses the grinding equipment, with connecting conveyors for ore feeding, also an air compressor, two vacuum pumps, bins for lead and zinc concentrate storage with individual conveying systems for loading the two kinds of concentrates. Filtering of concentrates is done on the second floor, the concentrates discharging directly into the concentrate-storage bins. On this floor are the lead and zinc filters, two diaphragm pumps, sample-filtering equipment, sample-preparation room and a change room for the mill employees. The flotation machines are on the third floor, at elevations allowing a gravity flow of the lead and zinc concentrates to the thickeners. The annex to the building contains a freight elevator and provides space for all reagent mixing and feeding, also storage of reagents and grinding balls. All the water used for milling purposes is pumped from the Junction mine and

Jan 1, 1947

By W. H. Emmons

MANY mineral veins occupy faults, and movements on certain warped fault planes have resulted in openings. On normal faults the [ ] steeper parts have the widest openings, and on reverse faults the flatter parts have the widest openings. Theoretically the width of the opening is the sine of the angle of deflection times the amount of movement on the fault plane, but actually the widths of some of the openings are probably less because of readjustments of walls along joints or of other readjustments. The first treatment of the subject seems to have been in a book on mineral veins by John Leithart, published in Newcastle, England, in 1838. Later the subject was treated in various texts published in Germany, in England, and elsewhere. The subject was not given much prominence in the earlier American literature on economic geology, but has interested several workers in that field in the past decade. In this paper examples of districts showing ore shoots on warped faults, with relations noted, are described. Fault veins of the three depth classes of veins are noted. Many epithermal veins are in normal faults, and commonly the larger ore shoots are in their steeper portions. Many mesothermal veins are in reverse faults and many others are in normal faults. The flatter parts of veins in reverse faults are commonly wider than the steeper parts, although they are not always as rich as the steeper segments of the veins. There are not many hypothermal veins in faults of measured displacement.

Jan 1, 1943

By Theodore B. Comstock

The existing topographical features of the United States present many points of interest to the student of dynamical geology, but there is, perhaps, no subject which offers a more promising field for investigation than the relations between the geological structure and the mineralogical characteristics of the principal mining districts. There are seven prominent culminating points in the relief of North America, each one of which can be used as a type of the physical features of all the others. Again, if we divide these seven elevated districts into two sets, we shall discover a still more striking parallelism between the individual members of each group. But the most remarkable revelation is the fact that each one of these points is more or less closely the centre of an area of economically important mineral resources. It is interesting to note in this connection that six of these mineral tracts lie mainly within the borders of the United States, three of them being almost wholly in our territory. For

Jan 1, 1883

By Leonard Obert

The structural stability of any mine or under- ground opening in rock is dependent on the stress field, that is, the state of stress in rock before mining, the stress distribution in the rock created by the opening, and the in situ strength of the rock. In the past mine design problems have been considered by assuming that the stress field is due only to the weight of the overlying rock, the stress around the opening can be calculated from elastic theory using laboratory-determined elastic constants, and the in situ strength of the rock can be determined from laboratory tests on small specimens. Solutions to hypothetical problems using this procedure are given in a report by Obert, Duvall, and Merrill.' Because of the uncertainties of these assumptions, this method can give only a rough approximation to the solution of actual structural problems. Obviously, this approach to determining the stability of underground openings would be greatly improved if the rock stress could be measured in situ, and it is undoubtedly for this reason that in recent years many investigators have spent considerable effort on the development of procedures for measuring rock stress. In this report a method is described for determining in situ the absolute stress in rock. The method requires the measurement of the deformation of a borehole before and after the rock is stress-relieved. The magnitude and direction of the applied stresses are calculated from the borehole deformation and modulus of elasticity of the rock. Also, the results of absolute stress determination made in the walls and pillars of underground openings in five rock types are presented. From these results a number of conclusions are drawn regarding the accuracy and limitations of this method and the problem of stress measurements in general.

Jan 8, 1962