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By C. P. Proctor, E. D. Hummer, J. A. Younkins

COMPARATIVELY little has been published relating practical experience with the control of the solid content of washery water systems that must be "closed." A "closed" system is one that for some reason may not be cleaned by the emission of any effluent of high solid content to be replaced by clean water in order to hold the solid content of the waste water within reasonable bounds for efficient washing. It is the purpose of this paper to describe practical experience with such a water system and to present data that it is hoped may be of some use to those who are faced with a similar problem. The washery clarification system described comprises three thickeners and three filters, and although it is admitted that each of the fields of water settling and thickening and filtration offers a wide scope for investigation, the authors are not attempting to present the answers to these problems but to set down the results of some experience along these lines, making no claims as to the ultimate solution of the problems in hand. The subject of water clarification has vast ramifications; for example, the science of flocculation as applied to water settling and filtering opens up fields of wide possibilities of which as yet little is known. A comparatively recent entry into the field of water treatment, which may offer possibilities of practical application, is the use of "wetting" agents such as "Aerosol." GENERAL ARRANGEMENT OF WATER SYSTEM So that the reader may obtain a comprehensive picture of the factors concerned in the control of a closed water system, it is necessary to give a brief description of the washing plant that is serviced by the water system to be described in the following pages. As we are not primarily concerned with the washing plant in detail, only a brief description of the flow of coal as it relates to the water system will be given here. For a more detailed account of this type of washer, the reader is referred to earlier papers.1.2 The washer, which handles 0 to 4-in. raw coal (750 tons per hour) is an American Rheolaveur plant of two complements; the coarse coal or sealed discharge plant and the fine coal or free discharge plant. The sealed discharge plant (Fig. r) consists of twin 48-in. primary launders, each having two washing boxes, supplemented by one 32-in. rewash launder, which also has two washing boxes. "Push" water for washing and conveying coal in these launders is supplied from a constant-head tank under pressure head of 10 ½ ft. at the point of introduction at the feed end of the launder. This tank also supplies water for upward or vertical currents used in the Rheolaveur or washing boxes. Each box discharges into a sealed boot elevator (Fig. I).

Jan 1, 1942

By William J. McCaughey, Richard L. Barrett

FOR sixty years-in fact, ever since the inception of the basic steelmaking process -basic slags have been the subject of study by chemists, metallurgists and petrographers, with the purpose of providing a better understanding of the reactions between the slag and the metal during the refining process. The result of this work and of the researches of mineralogists and physical chemists into the chemistry of oxide systems is a considerable body of literature dealing not only with slags themselves but with equilibrium relationships in systems synthesized from pure components. Since any actual slag is a multicomponent system more complex than any yet worked out in the laboratory, a complete solution to the problem is still far away. The writers have recently published a paper1 which they believe sheds light on the role of the slag in the elimination of phosphorus and it is the purpose of the present paper to discuss the metallurgical applications of that work. Since the pioneer investigations on the constitution of basic slags carried on by Stead and Risdale, Carnot and Richards, Hilgenstock and others in the 1880's, a considerable body of literature on this subject has appeared. Many of the more important of these papers were summarized in the previous paper,' therefore they will not be reviewed here. PETROGRAPHIC STUDY The present writers have had occasion, over a period of years, to examine petrographically a great number of basic steel- making slags having a very wide range of chemical compositions and made under various conditions of operating practice. Although certain of these slags exhibit some unusual features, it is a remarkable fact that the great majority of mature slags show certain typical characteristics, which persist throughout the wide range of chemical compositions. The typical mature slag usually appears under the microscope to consist of two phases. The phase of primary crystallization is light colored and doubly refracting. If separated mechanically and subjected to chemical analysis, it is found to consist essentially of SiO2, CaO, and P2O5. Sometimes, although not usually, two such silicate-phosphate phases are found. The other conspicuous phase is dark in color and optically isotropic, and occurs as the interstitial material between the grains of the light colored phase. This dark colored interstitial phase when analyzed is found to consist essentially of FeO,Fe2O3, MnO, and MgO. In some slags high in iron oxide, particularly those from low-carbon heats, dicalcium ferrite is found. This phase is dark reddish brown in color and anisotropic.

Jan 1, 1944

By E. S. Wheeler

The conversion plant of the Clirnax Molybdenum Co. is at Langeloth, Washington . County, Pennsylvania, approximately 30 miles west of Pittsburgh. The molybdenite concentrates converted originate in the company's mine and mill at Climax, Co1orado.l They are delivered to the Langeloth plant by rail packed in 175-lb. bags. War production of concentrates at the mill started in 1918 but it was not until about 1920 that any peacetime uses of the metal began to develop. Early conversion of Climax concentrates was mainly into ferromolybden um, produced for the company by a custom smelter. Chronology of Plant Development In 1924 a small hand-rabbled roaster was built at Langeloth, the roasted concentrates being used principally in the production of calcium molybdate by a conventional wet method. The size of the original plant is indicated by the fact that during 1925 about 400,000 lb. of molybdenum was converted. In 1925-1926, a multiple-hearth roaster was erected (12-ft. 0.d. 8-hearth Nichols-Herreshoff); also the first experimental ferromolybdenum was produced at Langeloth. At this time sa.1cium molybdate was produced in the furnace by a process to be described. With continued interest in and increased use of molybdenum, it was necessary to arrange for enlargement of plant facilities. Although only one 12-hearth 16-ft. 0.d. roaster was erected at this time (1929), plans were prepared for a duplicate unit, which was completed in 193 5, giving a potential roasting capacity of some 16,000 lb. Mo per day under the then prevailing roasting practice. Shortly after this second unit had gone into production, increased demand overtaxed the plant, and as a result two 12-hearth 18-ft. 0.d. roasters were placed in operation in 1936 and 1937. This gave the plant a potential roasting capacity of 40,000 to 50,000 lb. Mo per day, and necessitated major changes in auxiliary handling equipment as well as improved facilities for producing ferromolybdenum. Following the introduction (1938) of a new product—molybdic oxide briquets— and because of the constantly growing demand for molybdenum, plant facilities were again enlarged in 1940 and 1941 by the erection of two 16-hearth 18-ft. 0.d. roasters. These two new units, along with a change in roasting practice, brought the potential roasting capacity to 150,ooo to 190,000 Ib. Mo per day, based on about 15 lb. per sq. ft. of hearth area. Other departments of the plant have necessarily grown, until today the plant is modern throughout and completely mechanized, with equipment geared to produce marketable products as follows:

Jan 1, 1944

By John Birkinbine

IN the spring of 1878 the Pine Grove Furnace, located in Cumberland County, Pa., was blown in after lying idle for several years. The furnace was constructed in 1770, and for over a century it has been in almost continual operation. The plant consisted in 1877, when the writer was first called there, of a stone stack 32 feet in height, inclosing a shaft and boshes, the latter being 8 feet in diameter. The blast was supplied by two wooden blowing-tubs discharging into a third, having a floating piston sustaining a box weighted to give the desired pressure, the power being furnished by a waterwheel. A small 18-pipe hot-blast stove heated the blast. The furnace was remodelled during the winter of '77-78, and a Weimer blowing-engine with a blowing-tub 5 feet in diameter and 2 feet stroke, and the necessary boilers, were substituted for the wooden tubs and water-wheel. The stack was raised and enlarged, and provided with a bell and hopper having a central drop. Water dam and tymp were added, and general repairs were made. The hot oven, however, was riot increased. The furnace had always been operated with charcoal as fuel, and its reconstruction was made with a view to continue the use of charcoal, but provision was made for ample blowing capacity should other fuels at any time be employed. Although liberal arrangements had been made in cutting wood, the stock of charcoal supplied in 1878 was insufficient for the increased requirements of the remodelled plant, and notwithstanding the purchase of some 80,000 bushels from an adjacent idle furnace, there was riot enough to keep the furnace in constant blast until a new supply could be obtained, particularly as the season was unusually backward. Instead of following the established precedent of many charcoal furnace managers, i. e., blow out every February, it was determined to continue in blast, using coke as fuel. Accordingly, when the charcoal stock was exhausted, on March 22d, 1879, Connellsville coke was substituted, the change of fuel being made at once; that is, the coke charges followed immediately upon the last charcoal charges. After working a few days,-the strike in the Connellsville coke region cut off the supply, and anthracite coal was obtained, a mixture of the two fuels being employed. The strike continuing,

Jan 1, 1880

By Charles M. Behre

Under the caption "fault," geologists intend to include all mass movements of solid rocks over adjacent rock masses. When these are studied long after their origin, however, circumstances make it possible to recognize as "fault" only formations showing clear evidence of actual movement of one side with respect to the other. Necessarily the chief evidence consists of the offset of a vein, bed, or other key structure. Planes of movement parallel to the bedding, or nearly so, are recognized with difficulty or are overlooked and hence many important faults are wholly unrecorded. Indeed, among standard American references on structural geology, one alone emphasizes the importance of bedding-plane movements and that one only briefly1. For some time the writer, cognizant of this difficulty, has paid special attention to such movements, both as observed in the field and as recorded in the literature. It is believed that enough data are now at hand to justify a short discussion. Its purposes are fivefold. It is desired first to briefly classify and describe at least the more common kinds of bedding-plane faults, and second, to examine the causes and manner of their formation, at least in instances where the subject has not yet been well covered by other investigators. Third, certain by-products of bedding-plane movements are of such conspicuous importance to the economic geologist and mining engineer as to merit mention. Fourth, concrete illustrations should be placed before the reader in order to lend reality and to furnish a basis of comparison with his experience; such illustrations will necessarily be scattered through the text of this paper. They are only in part from the writer's own field observations. The fifth and prime objective of this paper is to direct attention to the great importance of movements on the bedding planes, in the hope that the recognition of such movements and their significance may simplify mining problems.

Jan 1, 1937

By John L. Boardman, C. W. Goodale

The motives behind the organization of the Bureau of Safety were twofold. First, there was the policy of the company toward its employees, which was one of fairness and consideration for their welfare and which entails the prevention of injuries; second, in order to prevent injuries, a study of their causes was essential, and this study could be made only by a thorough systematic classification of accidents, which required a special department to form and carry out the plans. The Bureau of Safety was organized to begin operation Jan. 1, 1914, hut practically all of the first year was spent in collecting data, making plans, etc. This department is semi-independent of others and reports directly to the president and general manager. It consists of the president of the company and the managers of the following departments: Mines, Washoe Reduction Works, Great Falls Reduction Works, Coal department, lumber department, B. A. & P. railway; and the chairman of the Bureau of Safety. Working under this group is the General Safety Committee consisting of the manager of mines, as chairman, the general superintendent, two assistant general superintendents, and the heads of the following departments: mechanical, geological, purchasing, sampling works, and framing plant; and also the safety engineer. Operating under the General Safety Committee are the group safety committees, consisting of the Anaconda group of mines, Boston & Montana group of mines, zinc group of mines, Washoe reduction works at Anaconda, B. & M. reduction works at Great Falls, coal department, lumber department and the R. A. & P. railway. Each of these group safety committces was in charge of the superintendent or manager of the group represented and the committee consisted of the foremen of the mine's or works that the committee served, as well as the safety engineer. These foremen and superintendents had met regularly for the purpose

Jan 1, 1923

By J. B. Austin, R. L. Rickett

Measurements of the rate of decomposition of austenite at constant temperature are commonly represented by plotting the percentage transformed on linear coordinates against time on either a linear or a logarithmic scale, a method that yields a curve that in shape resembles an integral sign (Fig. 1). A graph of this kind is not always convenient because extrapolation to the beginning or end of the transformation, or even intebrpolation, may be uncertain unless there are a large number of observations, which seldom happens; furthermore, the degree of consistency of a set of measurements is difficult to judge from such a curve. It is desirable therefore to find some mode of plotting, or even better to find a standard variety of graph paper, that gives a straight line instead of the integral shape. The straight line has many advantages; the number of readings required to fix the position of the curve is greatly reduced, interpolatiorr is easy and rapid, extrapolation can be carried out with greater confidence and, finally, the consistency of the measurements is readily estimated from the scatter of the points. The last advantage is particularly valuable when the data are obtained by direct microscopic examination alone, because they are influenced by the skill of the observer in estimating the relative size of areas, possibly also by undetected inhomogeneity of the specimen. A search for ways of representing the observations as a straight line yielded two methods which are described in this paper. The paper also includes an account of some interesting relations brought to light in applying these methods to the available experimental data. Usually the best way to derive such a method of plotting is to base it on established principles of chemistry and physics, as has been done, for example, for the temperature variation of the equilibrium constant of a chemical reaction. But so little is known of the mechanism of the decomposition of austenite, and so little is available in the way of well established background for changes of this kind, that it is simpler to rely upon purely empirical means. The general plan therefore is to study existing data with a view to finding some function of the direct observations that will

Jan 1, 1939

By R. R. Brandenthaler

CONTROLLING -the pressure under which a well produces crude oil has advantages that are ap¬parently not entirely realized by many operators in the Mid-Continent district. The possibilities of utilizing well pressures are indicated by a series of ex-periments recently completed by the Bureau of Mines in which pressure control was successfully applied to 20 wells in, the Davenport field. The flowing period of all of .the wells was lengthened in some instances it was. more than doubled. The daily oil production from approximately one-half of the tested wells was in-creased. The amount of gas produced per barrel of oil was decreased, and this, in addition to a slower rate of decline in production, indicated that the ultimate re-covery had been increased. The successful application of pressure control to flow-ing oil wells must take into consideration the relation-ships existing between pressures in the different parts of the system which, though not equal, are nevertheless more or less interrelated. These relationships, which have such an important bearing on the results obtained by applying pressure control, are as follows: A change in rock pressure will affect the tubing pressure, casing pressure, and differential pressure, but it will not affect the trap pressure. A change in tubing pressure will affect the differential pressure and casing pressure; also, a change in casing pressure will affect the tubing pressure and differential pressure. Trap pressures are not affected by rock pressure, casing pressure, tubing pressure, or differential pressure, unless the traps are required to handle gas and oil beyond their capacity. Trap pressures will affect the differential pressure, tubing pressure, and casing pressure and any changes at the well-head, whether differential pressure, tubing pres-sure or casing pressure, will affect the pressures in the well. Any change in rock pressure results in a change in the tubing pressure, casing pressure, and differential pressure. When any sudden or unusual change in well-head pres-sures occurs (for example, an erratic tubing pressure), and when a sudden decrease in oil or gas takes place, it is al-ways advisable to note whether the trap pressure has changed. If the trap pressure has not changed, note whether the oil line from the trap to the flow-tank has become clogged, as would be evidenced by a rise in the oil level in, the trap. If the oil line is functioning properly, and if the trap pressure has not changed, the flow-line from the well to the trap should be inspected. If it is free from obstructions, then changes in the differential pressure, tubing pressure, and casing pressure must be due to changed well conditions.

Jan 4, 1927

The Candelaria deposit was first discovered by Spanish prospectors in 1863; but its silver ore was not exploited on a large scale until 1873, when the Northern Belle mine was opened. The first Northern Belle mill was built at Belleville, 13 km (8 miles) to the northwest. From 1876 to 1886 the Northern Belle paid over $1 million in dividends each year. In 1882 a 43 km (27 miles) pipeline was built to Candelaria from the White Mountains to the west to furnish water for the mills and town at Candelaria; it was 100 mm (4 in.) in diameter, hand- riveted, and made in England. Sections of this pipe are used. as claim posts today. At its terminus at Metallic City, above Candelaria, a large concrete reservoir was built and called Metallic Lake. It is being used today by Candelaria Partners. Production at Candelaria declined after the mid-1980s, revived for a short period in the early 1890s, and during the 1920s a 300 ton mill operated for three years. Candelaria Partners is owned jointly by Occidental Minerals and Congdon and Carey, with Occidental being the operator. An engineering study done in mid-1979 resulted in engineering for the plant commencing in late that year and heap leaching being started in September 1980. Two million tons of ore per year containing 108 g Ag and 0.28 g Au per mt (3.15 oz Ag and 0.008 oz Au per st) will be mined from the Lucky Hill and Mount Diablo .areas about 1.6 km (1 mile) up Pickhandle gulch from the heap leaching operation. Waste -to -ore ratio is about 2:l. Ore is mined by front-end loaders and hauled by 85 ton trucks to the primary crusher, a 1,070 by 1,780 mm (42 by 70 in. ) gyratory. Primary crushed ore is conveyed to a stockpile of 10,000 tons live capacity from which it is reclaimed by hydraulic reciprocating feeders and conveyed to the crushing and agglomeration facility. This unit is composed of a 1.8 m by 2.6. m (6 ft by 12 ft) double deck screen, a 2.1 m (7 ft) standard cone crusher, and three conveyors. At the conveyor transfer points NaOH and NaCN solution is sprayed on the ore as it falls through a row of 75 mm (3 in.) hanging steel bars. During this process of turning and dividing the ore, the fines agglomerate with the coarse ore to give better percolation for the leach solution on the heaps. Provision was made for the later addition of 2 kg per mt (4 lb per st) of cement to aid in the agglomeration. Agglomerated ore at minus 50 mm (2 in.) size is then conveyed to a 390 mt (430 st) loadout bin where it is transferred to 85 ton rear dump trucks for haulage to the leach pads.

Jan 1, 1981

By L. D. Jaffe, D. C. Buffum, F. L. Carr, M. Goldman

IN an earlier investigation by Jaffe and Buffum,' a time-temperature transformation diagram for isothermal temper embrittlement of an SAE 3140 steel showed some suggestion of two overlapping "noses." These authors mentioned, as a possible explanation, that their specimens embrittled at 525°C were slightly softer than expected, perhaps because of some slight unknown variation in tempering treatment. Because a double nose would be of considerable importance in elucidating the mechanism of the temper brittleness reaction,' it appeared worthwhile to redetermine the portion of the diagram near 525°C. The same heat of SAE 3140 steel was used as in the previous work, and experimental procedures were the same.' Blanks were austenitized 1 hr at 900°C, water-quenched, tempered 1 hr at 675°C, and water-quenched. Groups were then embrittled isothermally at the temperatures and for the times shown in Table I, and water-quenched. For comparison, one series was tempered various times at 675°C, water-quenched, and given no further treatment. V-notched Charpy impact specimens machined from the blanks were tested over a range of temperatures. Energy absorption and percentage of fibrous fracture were plotted vs test temperature for each group, utilizing both the new data and those for the same treatment obtained in the earlier work,' except that the earlier data at 525°C were discarded.* are given in Table I. After a small correction (-0.28°C/BHN) for the variation of hardness" from Rockwell C24, the transition temperatures were plotted against embrittling time for each embrittling temperature, using the data of Table I for the range of treatments covered there, and using previous dataL for other treatments. By cross-plotting, Fig. 1 was finally obtained. It appears from this figure that if there are two noses to the time-temperature-transformation diagram for isothermal temper embrittlement of SAE 3140 steel, the upper nose must lie above 575°C. As interest had been primarily in the region near 525°C, a thorough study of embrittlement above 575°C was not made. Fig. 1 shows that such a study is needed, especially since preliminary measurements at the tempering temperature of 675°C reveal much more rapid embrittlement than would be expected from the data at lower temperatures. The transition temperature increased, approximately, as the logarithm of the tempering time (about 81/2°C, after correction for hardness, for each doubling of the time).& Since three tensile specimens tempered 240 hr at 675°C and water-quenched had an average yield-tensile ratio of 0.844, with 70.7 pct reduction of area, it is assumed1 that the tempering temperature is not above Ae1. The omission of a quench-and-reheat operation below Ae1, for the groups tempered at 675°C, as compared to groups embrittled at lower temperatures, should not affect results." More detailed investigation of embrittlement above 575°C is now under way.

Jan 1, 1954

By Reed W. Hyde

Some information has been published on the sulphur dioxide concentration of gases from D wight-Lloyd machines sintering lead ores but most of this relates to the customary practice in which the charge is relatively high in lead and is sintered in two passes Over the machine. It is thought, therefore, that the following observations on sintering in one pass a low-lead, low-sulphur charge may be of interest. The data were obtained in sintering lead sulphide notation concentrates mixed with minus 20.mesh oxide ore, the mixture carrying 16,8 per cent Pb and 5.3 per cent S. No returns were used because of the limited amount of sulphur available as fuel, but the mixture pelletized readily. to make a satisfactorily permeable bed on the sintering machine. As there was a possibility that the SO z might be utilized, a series of runs was made to determine the concentration ofSO2 in the gases at different points along the wind box, and the effect on this concentration of changes in the wind-box vacuum. Several runs each were made with vacuums (as measured in inches of water at the wind box) of (1 2 to 1 in., ('1 4 to 3% in. and (3) 6 to 5% in. The averages of the determinations from each series are platted On the graph (Fig, 1). The vertical scale shows the percentage by volume of SOz (as determined by the Reich method), and the horizontal scale shows the time e~pressed as percentage of the total sintering time. Sintering was considered to be complete when no SOz could be detected in the gases. In all cases the depth of bed was 4 1/2 in. and the weight of dry charge per square foot of hearth area was 45.2 lb. Other data are given in the table. The volume of gases corresponding to the several points along the wind box at which SO2 determinations were made could not be measured, but has been calculated from the SO2 Content and sulphur elimination on the assumption that the volume is uniform throughout the sintering period. This is not strictly true, as with a given vacuum the volume usually will drop slightly toward the middle of the period and increase toward the end. However, this made is a reasonable approximation. several interesting features are brought out by these data. In general, the published data on SO 2 concentration in sintering under ordinary conditions show a high SO2 concentration in the early stages followed by a rapid decrease in concentration and with relatively little SOz in the gases from the last half or third of the operation. In other words, the greater part of the sulphur elimination is obtained in the first third or half of the period. In these runs, the 6-in. vacuum, which approximates the usual operating condition, gives a curve that follows this customary pattern. Maximum SO2 concentration is obtained 234 min. after

Jan 1, 1944

By Paul A. Schafer

BaLatoc mine, the principal gold-producing mine in the Philippines, is 10 miles southeast of Baguio, in the Cordillera Central of northern Luzon.' The region is a deeply dissected plateau, ranging in altitude from 3000 to 5000 ft. above sea level. The rocks consist chiefly of granodiorite, diorite, and remnants of metamorphosed sediments, cut by dikes of andesite porphyry and quartz-diorite porphyry. All are intruded by a volcanic breccia plug, oval in horizontal section, about 3200 ft. long and 2300 ft. wide, which is the north central part of the mineralized area. Three main vein systems have been developed (Fig. I): 96, 231 and 364, which strike about N.70°E. and dip 50° to 90°S. Each system is a braided zone consisting of many individual components. They are disturbed by a series of premineral faults or shear zones striking about N.4s°E. Twelve of these produce important effects on the persistence, character, and quality of veins. From west to. east the faults are named (after claims): Hog, Big Gun, Bull, Pine, Lolita, Bar No. I, Bar No. 2, Curt, Sun, Helen, Itoe, and Frog. They are early breaks formed prior to the N.70°E. vein fractures. Mineralization occurs on them but, with several prominent exceptions, they are too low grade to mine. A number of northwesterly striking faults complete the structural pattern. Although they are relatively abundant, the effect on the veins is so small that they do not present any problems in mining or development. All of these faults that have been observed are later than the N.70°E. veins and offset them to the left. Northeast Faults The early northeast faults* are notably continuous in spite of their small displacements, 20 to 100 ft. Frequently they appear so weak that their importance is masked and their identity clouded. The gouge along any of these faults may vary from exceedingly weak at one point to very strong at another point only a short distance away. In other words, the amount of gouge depends on the changes of attitude of the fault plane, and is no indication of the extent of the movement. Places where walls are spread apart are filled with gouge squeezed from tight places. Also, mineralization sometimes obscures fault gouges by replacement. The mineralization of northeast faults has produced several important ore bodies; namely, 231 west on On fault and 96 vein on the same fault. To a smaller extent, Lolita, Curt, Bull, Bar NO. I, and Sun

Jan 1, 1949

By Will H. Coghill

In this investigation, the low-grade red iron ores of Alabama were examined by the use of "heavy solutions."' Ores are generally ground preparatory to concentration and as long as the concentrators make suitable products the effectiveness of the grinding is likely to be unquestioned; but when suitable products are difficult to make, an investigation becomes imperative. In such a case a dissection of the ore might be made to find the cause of the trouble. Such a condition of unsuitable products exists in the low-grade red iron ores of Alabama. They are difficult to concentrate. To be sure it might be said that the concentration of iron ores is a new thing and that the ore-dressing methods are undeveloped, but a glance at the industry reveals an abundance of methods and devices for almost all imlaginable conditions. Concentrating machinery has not been tested in this investigation. When the properties of the ores have been gleaned, the machinery may be used more effectively. Alabama Low-grade Red Iron Ores The Bureau of Mines has undertaken by laboratory methods to select the best means of concentrating the low-grade iron ores of Alabama. These ores constitute the parts of the ore bed left by selective mining. If more of the ore bed could be included in the current output, the mining cost would be reduced and longer life of the district would be insured. This advantage would be reflected by added permanence of churches, homes, schools and every undertaking in the commonwealth. The red ores chosen for this study of effectiveness of grinding are too low in lime to meet the requirements of the blast furnace. If lime is

Jan 1, 1927

By H. R. Fife

AN extended study of dustproofing Freeport-seam coal has empha-sized several fundamental features in the successful treating of bituminous coals. The extent of the exposed surface area and its absorption proper-ties for the dust-retaining film were found to be the most critical factors affecting the efficiency of fly-dust prevention. From recently published screen data and specific surface-area measure-ments on United Kingdom coals, an approximation of the surface exposed by mine-run coal and by 1-in. slack was made. The total surface area exposed by an average ton of this mine-run coal was estimated at 3768 sq. yd., while a ton of 1-in. slack represents a total surface area of 8779 sq. yd. with no additions for porosity. It is important in coal treating to consider that in mine-run coal the fraction above 1/4 in. in size represents 79.5 per cent of the total weight, but only 4 per cent of the total surface, while the fraction below 150 mesh contains 47 per cent of the total surface and only 1.2 per cent of the total weight. Also, one should appreciate that, in a 1-in. slack, the domestic stoker fraction (1 in. to 8 mesh) contains 76.3 per cent of the weight and 8.2 per cent of the surface, while the fraction below 150 mesh represents 3.3 per cent by weight and 54.8 per cent by sur-face area. Wax extraction data on 1-in. to 0 slack from mine-run treated Free-port-seam coal showed that the coating was distributed uniformly as a function of exposed surface; hence, nonuniformly on a weight basis. This distribution of the coating on mine-run coal therefore limits the extent of treating that is desirable on bituminous coal at this point in the preparation. The treatment of Freeport mine-run coal before it reached the tipple gave a 70 to 80 per cent reduction of fly-dust in the tipple, a satisfactory treated commercial slack coal, and an indicated change in the trend of tipple degradation. The premium-treated domestic sizes required a supplemental treatment for complete dustproofness.

Jan 1, 1937

By Harry M. Mikami

Chromite is the only ore mineral of metallic chromium and chromium compounds and chemicals. Because of this fact, chromite and chrome ore are used synonymously in trade literature. In commercial market quotations, chrome ore is the term commonly used. Chromite per se, because of properties imparted by its chromium content, is used in refractories and in special purpose moulding sands for metal casting. There are many other minerals containing some, generally minor, amounts of chromium, but none are commercial sources of the element (Thayer, 1956, p. 15). Chromium and chromite have many diverse uses that, directly and indirectly, critically affect vast segments of our modern industrial system. Most important are probably the metallurgical applications wherein chromium is a component of heat, abrasion, corrosion, and oxidation-resistant, and high strength alloys of many types. Stainless steels are the largest volume category of chrome-bearing alloys. Chromium chemicals are used in leather tanning, in pigments, dyes and mordants, printing, :hemica1 process industries, photography, pharmaceuticals, metal plating, pure chromium metal production, and for many other purposes. Chromite is a necessary constituent in basic refractories indispensable for the production of steel, copper, cement, and glass. Under present economic-political conditions, he U.S. and for that matter, North America, has virtually insignificant ore reserves of chromite. There are chromite deposits, as will be described, but they would only be used under highly abnormal conditions if all outside sources were cut off. All chromite used in the U.S. is imported from the eastern hemisphere in which are contained the principal world reserves. As the main suppliers in 1971-72 were the USSR, Turkey, Republic of South Africa, Philippines, and Rhodesia, the strategic nature of chromite is obvious.

Jan 1, 1975

By H. S. Rankin, E. C. Houston

OLIVINE is considered a valuable potential source of metallic magnesium in the chloride electrolytic process. Treatment of olivine with hydrochloric acid can be carried out under conditions that prevent the formation of gelatinous silicic acid, which would inhibit the reaction by forming a covering around the ore particles. It is our hypothesis that the formation of gelatinous silicic acid is avoided by the dehydrating action of the magnesium chloride, which takes up water to form a series of hydrates and leaves the silicic acid in granular condition. Leaching serves to separate the magnesium chloride from the silicic acid and unattacked ore, which can be treated with fresh acid in successive steps until extraction is essentially complete. The leach solution can be purified by the addition of caustic magnesia and the precipitated impurities removed by decantation or filtration. The resulting magnesium chloride contains: Fe, less than 0.001 per cent; Mn, 0.012 per cent; Ni, none. Chlorination[f] of olivine has two major defects: (1) a large part of the chlorine goes to form silicon and iron chlorides, (2) the anhydrous magnesium chloride is held by absorption in the finely divided silica. SUPPLY It is well known that olivine, a silicate of magnesium and iron, is present in large amounts in several localities in the United States. The largest deposits occur in the southeast and on the Pacific Coast.1 The southeastern deposits were surveyed by Lewis2 in 1896 and by Hunter3 in the period 1933-1940. It is estimated that high-grade olivine, conveniently located with regard to transportation and power, is present in excess of 230 million tons above local drainage level in western North Carolina and north Georgia. PREVIOUS TECHNICAL WORK In the past few years, olivine has attracted considerable attention in the refractory field. In the chemical processing industry, however, it has long been thought axiomatic that the cost of processing silicates is an economic barrier to their utilization, especially where more reactive minerals, such as carbonates, oxides or salts, are available. This attitude is not entirely justified in connection with olivine. Unlike most silicates, olivine is readily attacked by acids, chlorine and other reagents without resorting to fusion to break down the silicate structure. Olivine has been suggested as a raw material for making epsom salt by roasting with pyrites.4 Other processes include acid treatment5 and digestion with calcium chloride.6 Recently a small plant was built near Webster, N. C., for making epsom salt from olivine by treatment with sulphuric acid. This plant is now in commercial operation, producing approximately two tons per day. Olivine is also receiving attention from the fertilizer industry as an inexpensive source of soluble magnesium.

Jan 1, 1942

By J. M. Dahl, L. H. Van Vlack

STAFF: Editor, Gerhard Derge Associate Editor, Paul G. Shewman Carnegie lnstitute of Technology Schenley Park Pittsburgh, Pa. 15213 Editorial Assistant, M. A. Redmerski Production Editor, Otto T. Johnson THE METALLURGICAL SOCIETY: Jack H. Scoff, President; Richard C. Cole, Post-President; Harold B. Emerick, President-Elect; Ralph L. Hennebach, Treasurer; Robert W. Shearman, Secretary; Donald A. Parks, Assistant Secretary; Leslie S. Wilcoxson, Assistant Secretary. PUBLICATIONS COMMITTEE OF THE METALLURGICAL SOCIETY: R. E. Lund, Chairman; T. E. Dancy, H. B. Emerick, P. R. Gouwens, R. E. Grace, J. H. Jackson, H. H. Johnson, A. E. Lee, Jr., R. R. Nosh, R. D. Pehlke, and Paul Queneau. Review and selection of monuscripts is under the supervision of the following Publications Committees: Extractive Metollurgy Division: R. E. Grace, Chairman; G. M. Bell, D. W. Bridges, J. M. Cigon, G. W. Healy, T. A. Henrie, R. D. Holli-day, C. E. O'Neill, C. S. Sami s, A. W. Schlech-ten, and Charles Shull . Iron and Steel Division: R. D. Pehlke, Chairman; T. P. Floridis, Arnulf Muon, G. R. St. Pierre, E. T. Tuckdogon, L. H.Von Vlack, and J. M. Uys. Institute of Metals Division: H. H. Johnson, Jr., Chairman; J. B. Clark, Vice Chairman; M. R. Achter, G. F. Bolling, J. O. Brittain, G. P. Conard, II, L. A. Girifolco. G. T. Hahn, J.P. Hirth, L. Himmel, G. T. Horne, George Krauss, C. F. Leitton, Jr., J. D. Livingston, R. D. Monning, H. M. Otte, S. V. Radcliffe, R. E. Reed-Hill, H. C. Rogers, O. Sherby, M. T. Simnod, G. R. Speich, E. A. Steigerwold, R. A. Swolin, W. A. Tiller, W. R. Upthegrove, J. B. Wagner, Jr., Derek Walton, H. W. Weart, W. W. Webb, and D. WiIsdorf. Published monthly by the American lnstitute of Mining, Metollurgicol, and Petroleum Engineers, Inc., 345 East 47th Street, New York, N. Y. 10017. Telephone PLAZA 2-6800. Subscription $30 per year for non-AIME members in United States and North, South, and Central America; $32, foreign; $7.50 for AlME members. Single copies, $6.00, single copies foreign, $6.50. The AlME is not responsible for any statement mode or opinion expressed in its publications . . . Copyright 1965 by the American Institute of Mining, Metallurgical, and Petroleum Engineers, Inc. . . . Registered cable address, AlME New York . . . Indexed in Engineering Index, lndustrial Arts Index, and Chemical Abstracts . . Second class postage paid at New York, N. Y., and Ann Arbor, Mich. Volume 233, Number 1.

Jan 1, 1965

By A. W. Gauger

For perhaps one hundred years scientists have been engaged in attempts at devising a satisfactory method of classification of coal. During this time many charts, tables and graphs have been proposed, some of which have had considerable merit, but no one system has been found to be applicable to all ranks of coal, nor has any one system received general acceptance. In many of the proposals, water eliminated between room temperature and 105" C. has been regarded as extraneous matter and calculations have been made to the air-dry or water-free basis. Such methods of calculation do not lend themselves to satisfactory grouping of the high-moisture coals.1, 2 A consideration not only of the properties of the high-moisture coals and their behavior on drying but of the generally accepted ideas of the coalification process and the role of water in this process indicates that water is an essential constituent of all coals and therefore should not be neglected. For example, in such low-rank coals as the lignites of the Northern Great Plains' deposits the inherent moisture of the lignite as mined plays a very important part in the physical properties thereof. Freshly mined lignite contains upwards of 30 per cent water and is extremely tough and resistant to shock. Upon exposure to the atmosphere, the lump loses part of its moisture and checks and splits, leaving a friable and broken up residue. So far as we have been able to determine, water is the only thing lost in this process and we are forced to the conclusion that moisture plays an extremely important part in the structural properties of the original lignite. With reference to the coalification process, Mack and Hulett3 pointed out in 1917 the important part that water played therein. They observed that water cannot be removed from peat or coal by pressure alone, but only as the result of chemical reaction, and suggested that peat in its early stages may be largely a hydrosol. Slow chemical changes occur over a long period of time until finally the mass becomes more largely a hydrogel. These views are substantiated to some extent by the experi-

Jan 1, 1932

By Stanley Arnot

THE Mother Lode district in California is probably better known as the land of Bret Harte and Mark Twain than as a gold-producing district, although in this respect it holds an important place. The history of . quartz mining on the Mother Lode dates back to the '50s, when, after the first fever of the gold rush, men began to look for something more permanent, if not more lucrative, than the placers. Since that time,: the discoveries, shutdowns, and reopenings of the many mines, together, with the varying fortunes of their numerous operators, constitute an: interesting chapter in California history, but one too long to include in a description of mining methods. The district is well situated, so far as physical conditions are concerned. The Mother Lode extends along the western slope of the Sierra . Nevada Mountains, about half way between the crest and the Sacramento and San Joaquin Valleys, the country traversed consisting of low rolling hills-typical foot-hills. The elevation varies from 1000 to 1800 ft. and the climate is most desirable. There is an abundance of water, timber is easily obtained from the higher foot-hills a few miles east, and: cheap hydro-electric power is generated practically on the Lode. These natural conditions are conducive to low costs in mining and have been the deciding factor in allowing profitable exploitation of the comparatively low-grade orebodies. Transportation facilities are, not bad. Five branch railroads, from the main trunk lines in the valley, tap the district at intervals of from 10 to 30 miles, the roads connecting these are fairly good for trucking purposes. Labor is made up of several nationalities, the principal ones being American, Spanish, Jugo-Slav, Italian, and Mexican. There were few Mexicans on the Lode prior to 1919, but the shortage of labor, experienced generally by the mining camps of the country in the years following the war, made necessary the hiring of some. The percentage of Mexican labor rapidly increased in 1920 and for a year or so thereafter, but it has now diminished until at most of the mines only a few Mexicans are working. Of the different nationalities, Americans are the most efficient and Mexicans the least.

Jan 9, 1925

By John McCoy, Edward Thimons, Kenneth Heller

As the working levels in domestic mines are developed to greater depths and work areas move further from the shafts, maintaining a safe and healthful working environment becomes more difficult. Increased temperature and humidity at greater depths becomes first a factor in productivity and then a threat to the miner's safety. Initially, a miner might develop listlessness leading to reduced productivity and reduced alertness. Continued overexertion can lead to heat stroke and possibly death. Mine ventilation engineers, therefore, must not only provide clean air to work areas, but also must provide cooled air. The following article describes the United States Bureau of Mines (USBM) and Foster-Miller's successful development of a high performance, low maintenance direct water to air contact heat exchanger for mine cooling. While heat problems frequently do not develop until mining operations descend below sea level, problems will arise any time mining penetrates high temperature rock. Hot fissure water will increase the severity of the situation, causing increases in wet bulb temperature - the temperature to which perspiring miners reject metabolic body heat. All energy expended in mining adds to the heat load which must be absorbed by the ventilation air. Just when the cooling capacity of ventilation air is needed most, adiabatic compression (autocompression) of air descending the shaft raises air temperatures 1.3g°C (2.5'F) to 2.94"C (5.3OF)/305m (1000 ft) depending on humidity. Increased temperatures decrease the ability of the air to absorb heat from the mine, and work area temperatures rise. Once forced ventilation air can no longer maintain safe work area temperatures, refrigeration methods must be employed. Many mines now use cooling coils for this purpose. A cooling coil uses refrigerated mine water to cool the air by passing the air through banks of copper tubing filled with cold water. Heat from the air is transferred to the water through the pipe with the aid of fins. Conventional water to air heat exchangers prove troublesome for many mines be- cause of the high levels of dust that are encountered, causing these exchangers to operate at relatively low efficiencies. A more satisfactory approach to work- place cooling is desired - one that offers reduced maintenance and increased cooling efficiency.

Jan 1, 1982