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By H. S. Greismer

Alabama washes a larger percentage of its total coal output than any state in the Union. For producing coking coal, three-compartment jigs are favored; mines providing, steam and commercial coal use single-compartment jigs exclusively. Chemists are as necessary to washing plants as recording gages are to power plants. Experiments in Alabama indicate that the sludge released by coal-washing plants can be deposited in streams and on agricultural lands without disastrous results. CAMPBELL, 1 in 1896 said: "The Birmingham district in Alabama has certain great advantages for there are few places in the world where fuel and ore are so near together, although, unfortunately, both are of inferior quality; the ore being low in iron and high in phosphorus and the coal giving a weak and impure coke." This statement could not have been, questioned in 1894, but by 1896 several Robinson-Ramsay washers had been put into operation and weak and impure coke was a thing of the past. The Warrior coal field of Alabama contains six seams that yield coking coal of excellent quality. In descending- order, they are Brookwood, Milldale, Pratt, America, Mary Lee, and Black Creek. In each of these, however, bands of rock and bone coal of varying thickness are stratified with the coal; and while some of these impurities can be separated from the coal by the miner, much of it must be loaded out with the coal: Consequently, unless the coal is washed before being coked, it will yield a coke of varying ash and sulfur content quite unsatisfactory, for blast-furnace use. At present, the Gulf States Steel Co. is producing run-of-mine coal at its Virginia mine, from the America seam, that is uniformly low enough in ash to permit of coking without washing, but this is the only exception to the above statement. Alabama was the first state to experiment seriously with coal washers, was the first to install washers on a large scale and, up to the present, has maintained the lead. The Alabama State Mine Inspector reports that during 1923 20,919,303 tons of bituminous coal were mined in Alabama; 12,858,499 tons of which were delivered to coal washers for treatment.

Jan 9, 1924

By P. W. Rainier

The Chivor emerald field is situated on the eastern slope of the Andes in the Department of Boyach, at an elevation of about 8000 ft. above sea level. It overlooks the Llanos (plains) of the Orinoco and is watered by many streams flowing into the Rio Meta, the Orinoco's largest tributary. The Rio Meta is navigable from the Orinoco to a point within about 50 miles of the Chivor district and may eventually prove to be the reasonable outlet for the future development of this part of the country. The topography of the district is extremely rugged and mountainous (Fig. I), the difference in elevation between two points on a mining claim 1 km. square being sometimes as much as 4000 ft. The ground is covered with dense forest growth and the available timber is suitable and sufficient for mining purposes, but it is becoming gradually depleted by Indian clearings as the country around the mines becomes settled. The many streams combined with the rough topography give good possibilities for hydroelectric power. The center of population nearest to Chivor is Bogoth, the capital of the republic and a city of some 300,000 inhabitants; the distance is two and one-half days' travel. The first stage of the journey from BogotB is accomplished by rail to Choconta, some 4 hr. travel. From Choconth the journey is made in the saddle, one night being spent in Guateque. The latter is a small town and the nearest post office to the mines, being some 10 hr. ride distant. The roads are mere bridle paths, in parts hazardous and almost impassable in the wet season. All supplies brought to the mines, other than those produced locally, are carried by pack mule; mule freight from Choconth costs 3 centavos1 per pound. The climate of the Chivor district is healthful and bracing. The wet season begins in March and concludes in November, the remaining months being dry. The annual precipitation is about 130 in. The temperature throughout the year is nearly constant, varying from about 45" F. to about 65" F. shade temperature. The rains are cold and unpleasant, and for the greater part of each day in the wet season the country is shrouded in dense mist. A fire can with comfort be kept burning in a living room all the year round. There are a few slight

Jan 1, 1931

By D. A. Woodford

A number of observations have now been reported where the steady-state creep rate is apparently linearly dependent on the applied stress at high temperature and low stresses. This results in a creep rate greater than that predicted by extrapolation of high-stress data and has obvious practical implications. However, some of these observations are open to question. For example, the analysis by Harris and Partridge17 of the data of Greenfield et a1.18 was dependent on strain rates which, according to the original authors, were too low to be measured. Dr. Bernstein appears to have demonstrated a linear stress dependence at low stresses in two zirconium alloys. There are, however, a number of discrepancies in his treatment of the data and the subsequent theoretical analysis which require clarification. 1) It is incorrectly stated that steady-state conditions are revealed by a slope of unity on a plot of log i vs log (time). This objection is trivial if the author meant to say a slope of zero, although it is unclear how subsequent extrapolation was made to determine the steady-state creep rates at low stresses for the zircaloy-2. If the plot used was actually log c vs log (time), it is unlikely that a slope of unity would represent a constant creep rate. A plot of log (c - co) vs log (time), where c0 is the extrapolated intercept sf the steady-state creep rate on the strain axis, would have a slope of unity during steady-state creep. However, eo cannot be determined unless is is already known. 2) The author chooses to interpret his results based on the equation: where A, and B, are structure-sensitive parameters, Ql and Q, are the activation energies for the particular creep process, and n describes the sensitivity of the creep rate on stress. The first term represents a slip creep process and the second a diffusion creep process. However, it is stated later that "two distinct processes acting in parallel are involved". Eq. [I] surely represents two processes in series since each reacts to the total applied stress and t, is the sum of the individual creep rates. If the processes are acting in parallel, then the effective stress in each would be less than the applied stress. The component stresses would be given by sl + s2 = a but it is not clear how the magnitude of either, or the corresponding stress exponent, could be determined.

Jan 1, 1969

VITRIOL is likewise a mineral substance from whose exhalation, according to the opinion of some, are generated and produced the elemental materials or substances which produce metals, particularly gold. It is not a fumosity from metal nor is it a thing from which metal is extracted in any manner, although it shows in certain ways that it is not entirely without metal since it produces a certain trace of it. It is a thing that has a great resemblance to alum; it has a biting substance, is sharp to the taste and pungent to the tongue, and has an astringent nature. It dissolves in water and every damp place in a short time. They say that the reason for this lies, as with every other mineral, in its poorly joined watery substances. Nor does this mean that it does not contain much earthiness; on the contrary it seems to me that there are in it five. [2.9v] different participating qualities, that is, the property of sulphur, the action of alum, the corrosion of niter or of salt, and from the metals, the properties of copper and of iron. Its ore is mined in the valleys of mountains in wild but not too rugged places. It is more a gray earth or soft dead rock than a hard stone. It always has in it some yellow and green spots or near it is some sulphur mineral, either large or small.. When it is beneath the earth it exhales a sharp and very strong foul odor of scorching similar to that of sulphur. And the waters that rise where it is are all foul, earthy, and thick, some boiling with great fumes and others without great fumes, so that they surely appear to be infernal products in many places where there is a large quantity of it. After you have excavated and sorted well the quantity you wish, the ore must be put in a heap in open airy places before the vitriol can be extracted from this soil or ore, however you wish to call it. And it is left thus exposed to the weathering of the rains, the cold, and the sun for five or six months. It must be turned several times in order to bring to the surface that which was underneath; and always when this is done it is broken small with the eye of a mattock so that it is better heated and baked throughout. When it has been brought to this point it is covered over by a shed built above it or it is carried under a hut that has already been made. Here it is left to stand for another six or eight months before it is worked.

Jan 1, 1942

By Carl J. Christensen

This paper presents the development of diamond coring of the Weber sand section in the Rangely Field, Colorado. The description and operation of the diamond-coring equipment is included as well as the economics and results obtained by its use. Diamond coring is compared with the other methods used for drilling the Weber sand. Two types of diamond coring are discussed. The first and the one used most extensively is the regular 6 ?x 3¾-in. diamond bit and bottom hole 50-ft barrel. The second type, which has also been used successfully by Stanolind, consists of a 4 13/16 x 2 11/32-in. cutting bit and reverse circulation coring equipment. The first type will be referred to as regular or conventional diamond coring while the latter type will be referred to as reverse circulation diamond coring. Introduction Diamond coring has been used for many years in mines, quarries, surface coring at dam sites, laboratory work, and so on, but, until recently, had never been used very extensively in coring oil sands and other oil-bearing strata. It was demonstrated early in its development that hard forma-tions could be cut rapidly and that core recoveries far exceeded those obtained with conventional hard rock bits and quite often 100 pet recovery was obtained. However, the poorly constructed diamond-coring equipment and resulting high costs of early diamond coring slowed down its progress. Early in the development Of the Rangely Weber field, Rio Blanco County, Colorado, the Rangely Engineering Committee selected key wells that were to core the Weber section in its entirety. These cores were to be analyzed and various sections tested to gather sufficient information so that accurate reservoir studies could be made. The number of wells selected were kept at a minimum inasmuch as the productive Weber section, which is found at an approximate depth of 6000 ft, is thick, being as much as 600 to 700 ft on top of the structure. The Weber section is also comparatively hard to drill. For instance, one well took 60 days to core 490 ft of Weber sand with conventional coring equipment. It was noticed, however, that Core Labora-verse diamond plug cutter could quickly cut a plug from. one of these hard cores. This fact led to the serious consideration of diamond coring the Weber sand. The first diamond coring at Randy was performed by Stanolind in August 1946 andl although we experienced some difficulties, the results definitely indicated that diamond coring would be successful at Rangely for .coring the Weber sand. As more wells were cored and the equipment improved somewhat, diamond coring became less expensive than drilling the Weber sand with the conventional rock bit. In less than one Year diamond coring has definitely Proved its usefulness in oil-field drilling operations in the Rocky Mountain area where formations quite often drill very hard. The practice of spot coring in hard oil-bearing formations with conventional coring equipment is giving way diaimond coring of the entire sections, thus givng an operator the complete picture.

Jan 1, 1948

By Carl J. Christensen

This paper presents the development of diamond coring of the Weber sand section in the Rangely Field, Colorado. The description and operation of the diamond-coring equipment is included as well as the economics and results obtained by its use. Diamond coring is compared with the other methods used for drilling the Weber sand. Two types of diamond coring are discussed. The first and the one used most extensively is the regular 6 ?x 3¾-in. diamond bit and bottom hole 50-ft barrel. The second type, which has also been used successfully by Stanolind, consists of a 4 13/16 x 2 11/32-in. cutting bit and reverse circulation coring equipment. The first type will be referred to as regular or conventional diamond coring while the latter type will be referred to as reverse circulation diamond coring. Introduction Diamond coring has been used for many years in mines, quarries, surface coring at dam sites, laboratory work, and so on, but, until recently, had never been used very extensively in coring oil sands and other oil-bearing strata. It was demonstrated early in its development that hard forma-tions could be cut rapidly and that core recoveries far exceeded those obtained with conventional hard rock bits and quite often 100 pet recovery was obtained. However, the poorly constructed diamond-coring equipment and resulting high costs of early diamond coring slowed down its progress. Early in the development Of the Rangely Weber field, Rio Blanco County, Colorado, the Rangely Engineering Committee selected key wells that were to core the Weber section in its entirety. These cores were to be analyzed and various sections tested to gather sufficient information so that accurate reservoir studies could be made. The number of wells selected were kept at a minimum inasmuch as the productive Weber section, which is found at an approximate depth of 6000 ft, is thick, being as much as 600 to 700 ft on top of the structure. The Weber section is also comparatively hard to drill. For instance, one well took 60 days to core 490 ft of Weber sand with conventional coring equipment. It was noticed, however, that Core Labora-verse diamond plug cutter could quickly cut a plug from. one of these hard cores. This fact led to the serious consideration of diamond coring the Weber sand. The first diamond coring at Randy was performed by Stanolind in August 1946 andl although we experienced some difficulties, the results definitely indicated that diamond coring would be successful at Rangely for .coring the Weber sand. As more wells were cored and the equipment improved somewhat, diamond coring became less expensive than drilling the Weber sand with the conventional rock bit. In less than one Year diamond coring has definitely Proved its usefulness in oil-field drilling operations in the Rocky Mountain area where formations quite often drill very hard. The practice of spot coring in hard oil-bearing formations with conventional coring equipment is giving way diaimond coring of the entire sections, thus givng an operator the complete picture.

Jan 1, 1948

By R. Silverman, H. Blumenthal

It was found that despite the similarity of chemical analyses of different titanium carbides used as base materials for cermets, the physical properties, especially transverse-rupture strengths, of test bars were different. Hence this metallographic study attempts to link physical properties to micro-structures. It is shown that microstructure, grain shape, and grain growth are functions of three interrelated factors: 1—powder production procedure, 2—surface conditioning of the particles, and 3—impurities either contained in the original powder or acquired during ball milling. An explanation is offered for the "coring effect," long observed, but heretofore of unknown origin. The explanation is based on assumption of an oxide film and on chemical analyses which substantiate these findings. TITANIUM carbide has become in recent years a material of great interest in the high temperature field. Consequently, many manufacturers in the United States and Europe are producing titanium carbide for cermet applications as well as for additions to the well known tungsten carbide tools. All present commercial processes of titanium carbide production utilize the chemical reaction of titanium dioxide and carbon to form as nearly as possible stoichiometric Tic. This reaction is carried out in three ways: 1—in a menstruum of molten metal,' 2—in the solid state, either in a protective atmosphere2 or in vacuum;" or 3—in an are-melting operation. In spite of the fact that the pure carbides obtained in these operations are almost identical chemically, the physical properties vary considerably when they are combined with a binder (Ni, Co) to form cermets. This fact led the authors to examine metal-lographically nickel-bonded titanium carbide in order to find the possible reasons for this behavior. Materials and Methods Five different titanium carbides were used in this investigation. They are identified in Table I. The first four materials were used in the as-received condition. Material E, received in lumps, was crushed to —100 mesh and carried through a flotation process in order to bring its graphite content in line with the other products. A Galagher flotation cell was used with pine oil as frothing agent. The chemical analyses of the investigated materials are given in Table 11. The binder used was carbonyl nickel of 9 to 14 microns particle size, supplied by A. D. Mackay. The materials were ball milled at a ball to charge ratio of 6:1 using procedures described under "Experiments and Results." All particle sizes mentioned are averages determined with a Fisher Sub-sieve Sizer. Test bars (lx0.40x0.16 in.) were prepared by 1—hot pressing to 85 to 95 pct of theoretical density at pressures between 1 and 1½ tsi and temperatures from 1600" to 1800°C, 2-—-cold presssing after 3 pct camphor had been added, or 3—wet pressing, both 2 and 3 at pressures between 5 and 10 tsi. All pressed bars were sintered in a vacuum of 105 to 10-6 mm Hg for 2 hr at 1350 °C. Transverse-rupture strengths were determined by breaking on a Baldwin Universal Testing Machine over a 9/16 in. span. Densities were measured by water displacement. The preparation of the specimens for micrographs was done according to Silverman and Doshna Luscz." All magnifications are at X1000. A sodium picrate electrolytic etch was used. Experiments and Results The influence of ball-milling procedure, ball-milling medium, pressing procedure, and sintering procedure on the microstructure of 80/20 — TiC/Ni were investigated. Ball Milling of Materials A, B, and C in a Steel Mill: Figs. 1 and 2 show microstructures of hot-pressed and vacuum-sintered test bars of materials A and B after the respective materials had been ball milled to 2.1 microns particle size in a steel mill and mixed with 20 pct Ni binder. Material A (Fig. 1) shows considerable grain growth. Also evident is a tendency of the carbide grains to coalesce. The density is 98 pct and the low transverse-rupture strength of 111,000 psi is probably caused by many large grains and an unfavorable packing factor. Almost all grains show a slight indication of "coring." Material B (Fig. 2), although showing grain growth, still has many small particles and a better distribution of binder and carbide due to the relative absence of the coalescing tendency. "Coring" can be observed in almost all grains. The high transverse-rupture strength of 179,000 psi and the density of 100 pct are believed to be due to the many small grains completely surrounded by the binder phase. There is also a preference to form spherical grains with material A, while most grains of material B preserve their angular shapes. Material C, of which no picture is given, stays between A and B in every respect. Rounding of some grains can be observed as well as coring, but the latter to a lesser degree than with material B. Its densification is good and the transverse-rupture strength obtained is 142,000 psi. Ball Milling of Materials A, B, C, and E in a WC Mill: When the Tic powders were ball milled to 2 microns particle size in a we mill, then ball-mill mixed with 20 pct Ni binder, hot pressed, and vacuum

Jan 1, 1956

By L. A. Wilson, N. H. Harrison, R. R. Clay, R. L. Reed. D. W. Reed

Experiments have been performed in a linear near-adiabatic system for the purpose of extending data on reverse and forward combustion from atmospheric pressure to 1,000 psig. Results obtained from reverse combustion appear to conform qualitatively with the existing description of the process when reaction kinetics are suitably modified to account for the increase in pressure, i.e., increasing tbe pressure decreases the peak temperature and increases the combustion zone velocity. Forward combustion appears to be a fuel dominated process wherein peak temperature and combustion zone velocity are not very sensitive to changes in pressure. The moderate effects of pressure that do exist at low flux virtually disappear at high flux providing all oxygen is consumed. With this provision. increasing the pressure decreases the frontal velocity and increases the peak temperature. Results me shown graphicaIIy which demonstrate the effects of presswe on peak temperature, rate of advance, oil recovery, air/oil ratio, carbon oxides produced and temperature distributions. INTRODUCTION Numerous field tests of oil recovery using the technique of underground combustion are now in progress. Operating pressures used are always substantially in excess of atmospheric pressure. Nevertheless, the literature contains no laboratory data pertaining to the effects of pressure except for those of Martin, Alexander and Dew.1 Unfortunately, these data appear to reflect heat losses which may have obscured some of the effects of increased pressure. The underground combustion processes are exceedingly complex, and general concepts relating to them have necessarily been described in relatively simple terms. Particularly with regard to forward combustion, the interaction between multiphase fluid flow, heat transport and various rate mechanisms is so involved that to date a rigorous treatment is not available. In this paper a linear near-adiabatic system is described which was found capable of operating at pressures up to 1,000 psig and temperatures of at least 1,100 F. Using this equipment, experiments on both forward and reverse combustion were accomplished at elevated pressures and the data presented as a function of air flux or temperature. It will be assumed that the reader is familiar with the papers of Wilson, Wygal, Reed, Gergins and Henderson on forward combustion and that of Reed, Reed and Tracht3 on reverse combustion, so that it is not necessary to repeat the details of the two processes. EQUIPMENT AND PROCEDURE It was decided to study the effects of pressure over the interval 0-1,000 psig using a near-adiabatic system. Since satisfactory operation of such a system requires very thin tubing walls, it proved necessary to insert the entire combustion tube and heater assembly into a high pressure jacket in such a way that the internal tube pressure could be counterbalanced by pressurizing the annulus. This meant that the numerous heater and thermocouple wires had to pass through gas-tight seals that required considerable effort to maintain effective. An unexpected difficulty arose in connection with the nature of the annulus pressuring gas. It turned out that the composition of this gas critically affected the functioning of the adiabatic control system. Operation of the equipment at elevated pressures was not particularly difficult. The problems that did arise were a direct consequence of the large numbers of connections and seals involved. EQUIPMENT The equipment consisted of the combustion tube which fitted within the pressure jacket. This jacket was in turn mounted on a wheeled carriage which facilitated movement and provided for rotation of the

By L. E. Mannion

Sodium carbonate or soda ash is one of the two principal alkalis of commerce, the other being sodium hydroxide. The use of sodium carbonate is recorded in ancient Egypt, where naturally occurring brines or solid salts were the source of impure soda for glass manufacture. In early modern times until the 19th century, soda ash, as well as potash, was recovered mainly by leaching the ashes of plants. In Europe, during the 18th century, a brown lye for laundering was produced by pouring hot water over burned seaweed. One of the better grades of this kind of material was a product from Spain termed "barilla," which analyzed 24 to 30% Na2CO3 (Hou, 1942). The birth of the modern chemical industry may be said to start with the LeBlanc process for making sodium carbonate from salt, sulfuric acid, and lime. Although invented in France in 1691, England made the most extensive use of it. In 1863 Ernst Solvay devised the ammonia-soda or Solvay process, and by 1874 the first of many Solvay plants was built. These were to supplant the LeBlanc method and dominate soda-ash production until the present time. Although of lesser importance, natural sodium carbonate deposits continued to support operations in various parts of the world. Shortages during World War II stimulated production of "natural" soda ash, particularly at Searles Lake, Calif., and at Lake Magadi, Kenya. Within the last ten years in the United States, a notable surge in production has occurred based on the vast trona deposits in Wyoming. Table 1 lists some principal producers of natural soda ash and the character of the source deposits. Uses and Specifications Sodium carbonate (Na2CO3) is a white, crystalline hygroscopic powder, an aqueous solution of which is strongly alkaline. Most commercial soda ash is a product that must meet high standards of purity and uniformity. Two forms are commonly sold: a fine, powdery material called light ash, and a more coarsely crystalline dense ash; these weigh 32 to 39 and 60 to 66 lb per cu ft, respectively. Dense ash may be produced by adding a little calcium chloride to the sodium carbonate solution to control crystal size and prevent agglomeration. Specifications and typical analyses of light and dense ash in the United States are shown in [Table 2]. Sizing generally achieves about 75% of light ash and 90% of dense ash between +30 and -100 U.S. screen mesh. World use of sodium carbonate now exceeds 24 million tpy, of which the United States produces more than 7 million tons. The market for soda ash is dominated by the glass industry, which uses almost half the supply. Inorganic chemicals, including phosphates and silicates, consume about 30%. Other principal uses include the manufacture of soaps and detergents, petroleum, caustic soda, organic chemicals, and paper; and nonferrous and ferrous metallurgy. [Fig. 1] shows the percent consumption by use of soda ash in Europe, Japan, and the United States. The price of soda ash in the United States remained relatively stable during the 1960s at about $33 per ton; in 1972 it reached $35 per ton. The price in 1975 was $42 per ton. Geology Mineralogy [Table 3] is a list of the more common sodium carbonate-bearing minerals. Anhydrous sodium carbonate is almost never found in nature; only natron, thermonatrite, and trona have been used directly as sources of sodium carbonate. At Searles Lake, burkeite and hanksite are important in the reconstitution of

Jan 1, 1975

By N. S. Greensfelder

THE firing of explosive charges by electricity dates back to 1745 when a Doctor Watson is said to have used an electric spark for igniting gunpowder. His method failed in practical application because, to ignite blasting powder, heat must be applied for a period longer than the duration of an electrostatic spark; Wheatstone found that this does not exceed 14000 sec. The usual result, therefore, of an electrical discharge in gunpowder is to scatter the grains, without igniting them. Benjamin Franklin partly overcame this difficulty by confining the powder and crushing grains that lay in the line of discharge. In 1822, after the primary cell had been discovered by Volta, Robert Hare, professor of chemistry at the University of Philadelphia, noticed the heating of a short piece of wire placed in the circuit of a voltaic battery. Later work by Oersted, Faraday, Wheatstone, and Henley resulted in the development of the magneto blasting machines. Fulminates of gold and silver are thought to have been discovered by Cornelius Drebbel, about 1600. Pepys, in his diary for November 11, 1663, recounts a conversation with Doctor Allen, who told him about aurium fulminans "of which a grain ... . put in a silver spoon and fired, will give a blow like a musquett and strike a hole through the silver spoon downward, without the least force upward.1," This fallacious belief that explosives act downwards only, is still held by many who should know better. Fulminate of mercury was discovered by Howard in 1799. J. Shaw of Philadelphia is said to have invented a fulminate-of-mercury cap in 1814; his idea was later adopted by gunmakers in England and France. The principal advantages of

Jan 9, 1922

By R. T. Bayliss

Although every mill-man of even limited experience in the amalgamation of gold-ores is probably aware that copper plates will in time become coated by the accumulation of gold-amalgam, it may be that many do not know to what extent this accumulation occurs in the treatment of ores dissimilar to those with which they are familiar, and under conditions foreign to those with which they have to contend in any given locality. The writer, therefore, feels relieved from the necessity of offering an apology for dealing with such an elementary subject, believing that certain facts in reference thereto, which have recently come under his personal observation, will be interesting to many, and possibly instructive to some persons engaged in the amalgamation of gold-ores. The facts hereinafter stated were observed in the 50-stamp "combination"-mill, owned by the Montana Mining Company, Limited, of England, operating the Drum Lummon mine, situated at Marysville, Montana. The material treated in this mill consists of quartz and ores of somewhat variable character, containing gold and silver in native form, as well as in chemical union with the sulphides of iron and copper. There is also present a little lead, some arsenic and antimony, and a little zinc. The last fair minerals, however, are of irregular occurrence, and but sparingly distributed throughout the ore, and yield only a trace in the analysis of the average material in this mill. The free gold, which is in a condition of fine division, and rarely visible to the naked eye, carries silver as an alloy; and the native silver, which is only occasionally observed, is present in the richer ores in wireand in leaf-form. Of the two precious metals, the ore worked during the past three years and ten months, the period under review, contained about 1/2-ounce of gold, and from 7 to 12 ounces of silver, per ton of 2000 pounds.

Jan 1, 1897

By Edgar Newhouse

THE lead smeltery at East Helena, Mont., controlled by the American Smelting & Refining Co. since 1899, has been in continuous operation for the past 25 years. Most of the old smelting and roasting practices have been superseded by newer and more modern methods, but on account of the difficulties involved in changing the original construction, some of the methods in use of handling materials are not the most modern. This plant, however, is particularly interesting as being one of the few essentially lead smelteries in the United States. At Federal, Ill., Collinsville, Herculaneum, and possibly other places, similar work is being done, but it is doubtful whether any of them produces as' much lead as the East Helena plant, whose monthly production averages 6,500 tons. The situation, 5 miles east of Helena, is convenient to both the Great Northern and Northern Pacific railroads, while the electric power is obtained from Canyon Ferry, 12 miles east on the Missouri river. Local quarries supply the requisite amount of lime rock, Which the contracts require to be crushed to 5 in. Of the 450 men on the monthly pay roll, a census showed 200 Austrians, 110 Americans, and 30 Italians, the balance being made up of miscellaneous nationalities. The ore supply is mainly derived from the Coeur d'Alêne mines in Idaho, and consists of crude ore, coarse and fine concentrates, middlings, and slimes. About one-half of the total output of the Coeur d'Alêne camp is sent to East Helena. An average analysis of these ores gives approximately 40 per cent. lead and 11 per cent. sulphur. Local shipments, varying, from month to month, are mostly highly siliceous gold and silver ores carrying little lead. Small shippers, leasers of old mines, and prospectors are very numerous throughout Montana, and it is the policy of the company to encourage these men by every possible means, including the free sampling and analysis of their ores and aid in the development of their mines and prospects whenever their results show promise.

Jan 8, 1914

By Douglas Bunting

THE first presentation of this paper was before the Pennsylvania Anthracite Section of the Institute in May, 1914, after which a committee was selected to verify and add to the data contained in the original paper. The author is indebted to G. W. Engel, H. W. Montz, R. V. Norris, and H. H. Otto, of this committee, for additional information. I also wish to acknowledge the courtesies extended me by a number of the mining companies for information relative to accidents, systems of proving, and mining practice. The paper should bring forth additional criticism and recommendations, which will be of value in establishing future practice relative to this subject. Throughout the northern anthracite field of Pennsylvania glacial and alluvial deposits of sand, gravel, and clay exist in varying quantities. The extent and character of these deposits are described by N. H. Darton in Bulletin 45 of the U. S. Bureau of Mines, entitled, Sand Available for Filling Mine Workings in the Northern Anthracite Basin. It is estimated that the northern field contains 10,000,000,000 tons of these deposits, or a sufficient quantity to cover the 176 square miles of coal area to a depth of about 25 ft. These deposits of sand, etc., overlying the coal measures attain a maximum depth of more than 300 ft., and are usually saturated to such an extent as to render them semi-fluid. It is .this condition of fluidity that limits the mining of coal seams cropping in or in close proximity to these deposits, and it is in the interest of safer and more efficient mining methods under these conditions that the study of this -subject is here presented. The dangers incident to mining under deep deposits of sand have long been realized by operators in the northern anthracite field, and much has been done to prove the contour of the rock and vein croppings underlying these water-bearing deposits of sand and gravel.

Jan 1, 1915

By E. Robert Ruhlman

The term potash refers to potassium oxide (K2O), a compound not found in nature or produced by man but used as a basis for comparison of all potassium compounds and now is applied generally to various compounds of potassium used by agriculture and industry. Potassium, the last of the three major plant foods to be used in appreciable quantities by agriculture, was originally produced only from organic materials for chemical uses. The first recorded identification of a mineral source of potassium was in 1797; but, until the discovery in 1839 of the large deposits of soluble potassium salts in Germany and their development in 1860-61, nearly all potassium compounds continued to come from organic sources. With the importance of potash to plant life discovered in 1840 by von Liebig and the production of potash in Germany after 1861, agriculture soon became the major consumer of potash. In 1958 agriculture accounted for about 95 pct of the total potash consumption; the chemical industry used the remaining 5 pct. Mineralogy and Geology Potassium, an important constituent of the earth's crust, occurs in many forms in numerous rocks and minerals, including water solutions, but is not found in its native state. Twenty-five of the more important potash minerals are listed in Table 1. The properties of the five potash minerals most common in Permian salt formations follow: Sylvite is isometric; has a hardness of 2, a specific gravity of 1.98, vitreous luster, and cubic cleavage; and ranges in color from clear to brick red. In taste it is similar to halite but has a sharper taste and a cooling effect on the tongue. It is very plastic under pressure. Langbeinite is isometric; has a hardness of 4, a specific gravity of 2.82, vitreous luster, conchoidal fracture, and no cleavage; and ranges in color from clear to gray. It is taste- [ ] less, somewhat granular, and slowly soluble in water. Polyhalite is triclinic; has a hardness of 3, a specific gravity of 2.77, and dull luster; and

Jan 1, 1960

By R. Ellerman

The hydrometallurgical processing of an impure zinc oxide fume is described. Flowsheet includes roasting for fluorine elimination and countercurrent leaching to produce a neutral sulphate solution. This solution is combined with zinc sulphide leach solution for subsequent purification and electrolysis. Novel to the neutral leach step is the use of automatic pH control. THE electrolytic zinc plant of the Hudson Bay Mining and Smelting Co. Ltd. was erected in 1929-1930 to process zinc sulphide concentrates produced from ore mined at Flin Flon, Manitoba. Since 1930, production of high purity slab zinc has been continuous. In May 1951, construction of a fuming plant was completed and operations were started to recover zinc from copper smelter reverberatory slag. A second zinc-bearing material in the form of an impure zinc oxide fume was then made available for treatment. Coincident with the construction of the fuming plant, additions were made to the electrolytic zinc plant to recover the zinc in the fume as slab metal. The original plant capacity of 70 tons of slab zinc per day was steadily increased up to May 1951, at which time the daily production averaged 130 tons. With the advent of the fuming plant and the processing of zinc oxide fume, production was stepped up to the present capacity of 190 tons per day. Of this, approximately 48 pct is contributed by the processing of zinc oxide fume. This additional source of raw material has allowed a 20 pct cutback in mining and milling operations, thereby adding considerably to the life of the Flin Flon mine. Zinc sulphide concentrates are treated by the usual hydrometallurgical method. Details of this operation and plant were described by Carr and Reikie in 1935' and Kent in 1951.8 The impure fume is treated in two operations. First, it is roasted to eliminate fluorine. Next, it is subjected to a countercurrent leach with spent electrolyte to extract the zinc as a neutral zinc sulphate solution. During this leaching step, soluble impur-

Jan 1, 1956

By Ottar Dragge, C. Danielsson, Bo Kalling

THE desulphurizing of pig iron has been accomplished with a number of different additions. The oldest and still most commonly used agent is soda, the extensive use of which commenced about 1925, when it was used principally for cupola furnace iron. More recent experience' seems to show that better results can be obtained with sodium hydroxide. The well-known desulphurizing properties of lime have also been exploited in different technical processes. Another material with even more powerful effect is calcium carbide.' The desulphurizing ability of manganese, when added to the ladle in sufficient quantity, should also be mentioned in this connection. During recent years increasing attention has been paid to the desulphurizing properties of metallic magnesium." An addition of a suitable alloy of magnesium is now in use purely for the purpose of sulphur elimination. Of the desulphurizing agents mentioned, lime is by far the cheapest, provided that the reaction can be brought about rapidly and completely. Therefore, a method that makes full use of the desulphurizing ability of lime may be able to compete with other processes. A method developed at the Dom-narfvet Iron and Steel Works (Sweden) will be described, which enables pig iron to be rapidly desulphurized to very low sulphur contents by using a burnt lime powder. as the desulphurizing agent. Lime in Older Processes In cases where lime has been used for the desul-phurization of pig iron, it has generally not been used alone, but mixed with other substances such as fluorspar, to obtain the formation of a molten slag during the process. This method has been tried by Tigerschiold,' who treated the iron with a lime-fluorspar mixture, the stirring of the iron being brought about inductively with low frequency alternating current. Very good results were obtained. A process of this type has also been suggested by R. P. Heuer, U. S. A. The principles of this method, which has been tested in Great Britain by Newell. Lanener. and Parsons." re that a mixture of lime and fluoispar is added to the hot metal in the ladle, while a powerful stream of nitrogen gas is blown into the bath to produce the required intermixing. The results of the tests were unsatisfactory, however. A similar process has been developed at The Steel Co. of Canada, according to a statement by H. M. Griffith.' Here the tests were carried out in a carbon-lined ladle provided with carbon tuyeres in the side wall for blowing nitrogen into the bath. The addition consisted of about 20 lb of a mixture of burnt lime and fluorspar per ton of pig iron. Good results appear to have been achieved. The sulphur content of the pig iron is stated to have been reduced from 0.025 to 0.050 pct down to 0.006 pct. Various methods of desulphurizing pig iron have been tried using lime powder without fluxing material for fusing. Eichholz and Behrendt7 have experimented with blowing a powdered limeicoke mixture with air into the ladle. Their results were, however, not conclusive and the experiments do not appear to have been continued. Similar experiments have been carried out at Domnarfvet, using nitrogen instead of air in order to avoid oxidation. But these attempts were not particularly successful. It appears to be difficult to achieve the required agitation by this means. The strong cooling effect of the gas on the iron is also a serious drawback. A method in many respects similar to that tried at Domnarfvet was tried by Eulenberg and Krus at the end of the 1930's. Here again desulphurization was carried out with lime alone, brought into contact with the molten iron in a rotary furnace. The temperature was kept at the required level, 1400" to 1500°C, by the introduction of a pulverized coal burner in one end of the furnace. The speed of rotating was not given. A paper by Bading and Krus states that, in one of the first experiments, the sulphur content in 56 tons of pig iron was brought down from 0.186 to 0.035 pct in 117 min, but that a considerable shortening of the time would be possible. According to later reports by Eichholz and Behrendt,' it should be possible by this process to achieve a desulphurization speed of 0.35 pct S per hr for a consumption of 6 to 10 pct limestone and 2 to 3 pct coke, as fuel exclusively. The final sulphur content is, however, not stated. Domnarfvet Method After a number of different procedures had been investigated, the tests at Domnarfvet were directed to desulphurization with lime in a rotary furnace. Before going into the practical details of the method, the theoretical aspects will be discussed briefly. If the pig iron does not contain alloying elements other than carbon, the reaction can be expressed most simply by the usual equation: FeS + CaO + C = Fe + CaS + CO [I] 4H,. ~ 34,000 cal That this reaction can be carried through to a very complete desulphurization of pig iron has been shown by OelsenD in a discussion in connection with the Eulenberg and Krus' method. He mentions two laboratory tests, in one of which the sulphur content in the pig iron at 1400°C was reduced from 0.540 to 0.006 pct after treating with 3.35 pct lime. The pig iron had a low manganese content, but other analysis is th. given. Mention also should be made of the recently published investigations by Fischer and Cohnen'" dealing with the influence of the carbon content of the iron on desulphurization with lime, although in this case fluorspar was added also. The tests show that efficient desulphurization is possible with lime in the steel bath, provided that the carbon content is sufficiently high. The temperature employed in these tests was considerably higher (1620") than that normal for treatment of pig iron. The author concludes that the product S% X C% - 0.011 at the temperature in question.

Jan 1, 1952

By C. Wagner

When an alloy solidifies and the composition of the solid differs from the composition of the liquid, atoms of the alloying elements rejected at the solid-liquid interface have to diffuse toward the bulk liquid. Diffusion may be supported by convection. Theoretical calculations have been made for a liquid without convection, for a liquid involving natural convection, and for solidification of a liquid alloy at the surface of a rotating disk as an example of forced convection. WHEN an alloy solidifies, the equilibrium composition of the originating solid phase differs in general from the composition of the liquid phase. Thus, segregation may take place, i.e., the composition of the solid phase formed at different times differs from the bulk composition. The degree of segregation depends on several factors such as the partition ratio of the alloying elements between the liquid and the solid phase, the rate of solidification, and convection. To clarify, a theoretical analysis has been worked out. The following presuppositions are made: 1—Calculations are confined to the solidification of binary alloys in which the concentration of component A (solvent) is much greater than that of component B (solute). 2—The equilibrium concentration of the alloying element B is assumed to be lower in the solid phase than in the coexisting liquid phase. 3—At the solid-liquid interface, partition equilibrium is assumed. 4—The dependence of the partition ratio of the solute between the liquid and the solid phase on concentration and temperature is disregarded since only small concentrations of the solute are considered. 5—The concentration of the alloying element B in the solid phase is supposed to be lower than the saturation concentration for the formation of a second solid phase. 6—diffusion in the solid phase is disregarded, since diffusion rates in the solid state are comparatively low. When the partition ratio of the solute between the liquid and the solid phase is greater than unity, a part of the solute is rejected at the solid-liquid inter- face and has to diffuse toward the bulk liquid. Thus there is necessarily a concentration difference in the liquid between the solid-liquid interface and the bulk liquid as has been pointed out by Hayes and Chipman,' Rutter and Chalmers,' Tiller, Jackson, Rutter, and Chalmers, and others (see Fig. 1). In view of the concentration gradient in the liquid phase, the ratio of the concentration of the solute in the solid phase to that in the bulk liquid differs from the equilibrium ratio, although at the solid-liquid interface, partition equilibrium is assumed. Under certain limiting conditions, the solid may even have virtually the composition of the bulk liquid as is shown below. General Equations Let u be the linear crystallization velocity in cm per sec, D the diffusion coefficient of component B in the liquid phase in cm2 sec-', c',, and c", the concentrations of component B in the liquid and the solid phase, respectively, in mol per cm", and s the distance from the solid-liquid interface toward the bulk liquid at a given time t. The partition ratio is assumed to be equal to a partition constant K. Thus. The difference in the specific volume between the liquid and the solid phase may be disregarded. In view of the discontinuity of the concentration of the solute at the interface according to Eq. 1, u(c', — c",), , , = tic', (S = 0)(1 — K) gram-atoms B per unit area per unit time are rejected at the interface and have to diffuse toward the bulk liquid. From Fick's first law it follows that uc',,(s = 0) (1 ~K) == - D(dc',,/?>s). _-„. [2] The concentration profile in the liquid near the interface depends on the interplay of diffusion and convection as is discussed in the following sections. A schematic concentration profile is shown in Fig. 1. The "effective thickness of the diffusion boundary

Jan 1, 1955

By C. G. McLachlan

This paper outlines a series of tests carried out during the first six months of 1927, in the laboratory and mill of the Moctezuma Copper Co. The object of the tests was to determine whether the soluble salts present in the mill circuit water were having a detrimental cffect on the recovery made by flotation. To accomplish this object a series of water analyses were made, followed by flotation experiments in which definite quantities of the salts found were added as addition agents. Before proceeding to describe the results obtained in the course of the investigation, reference will be made to five articles that have appeared during the last two years in which the action of such salts is discussed. Two of these are by A. T. Tye1, 2 and describe flotation practice at Cananea and the way in which water high in ferrous and ferric salts is conditioned for differential flotation. Two, one by A. C. Munro and H. A. Pearse3 and the other by H. A. Pearse,4 outline Britannia milling methods, and the advantages that have been derived from ore washing; while in the fifth, A. W. Hahn5 summarizes the results of laboratory experiments in which ferrous, ferric and aluminum sulfates have been added in varying amounts. The mill of the Moctezuma Copper Co. has a daily capacity of 3000 tons, and is situated about 75 miles south of Douglas, Arizona, on the Rio de Nacozari, in the Mexican state of Sonora. The ore treated is mined from the company's mine at Pilares, about 5 miles southeast of the Concentrator. This ore contains about 2.5 per cent. copper— present almost entirely as chalcopyrite—and 10 per cent. iron, of which

Jan 1, 1928

By H. T. Matson, G. Allan Wallis

In the Northern Rhodesia copper fields the size of the orebodies and the exceptionally consistent values over great distances made it possible to outline the ore with drill holes spaced at 1000-ft. intervals; in some cases, 2000-ft. intervals. Owing to the fact that the size of the deposits made it very evident that much outside capital would have to be called in to develop them and, owing to the wide spacing of the holes, it was essential that the sampling of drill holes be as accurate as possible, so that outside examining engineers could adopt the results as trustworthy within the limits of accuracy of the drilling method. With this object in view, the field staffs of the Selection Trust group of companies have studied for the past three years, and are still studying, drill sampling in a very thorough manner, and for this reason it is felt that these notes may be of interest to engineers or students who are, or in the future may be, in charge of drilling operations. The orebodies in this field lie in shales and sandstones, more or less indurated, in the lower strata of the Roan series, and the copper mineralization is in the form of sulfides predominantly, with some oxidized copper minerals. These interested in the geology and history of the area can find full details in a recent paper by Anton Gray and Russell J. Parker.' In some formations core recovery was good—between 90 and 100 per cent.—and where any difficulty occurred in recovering sludge, as in fissured ground, this was not insisted upon; but the samplers' difficulties, both practical and theoretical, do not commence until core recovery drops and sludge recovery becomes important, for if core is being ground up it is recoverable only as sludge, and naturally the sludge sample, which theoretically should contain only the material cut from the annular ring around the core, is being salted, in a sense, by the ground-up core. In the perfect core, where 100 per cent. core is recovered as well as 100 per cent. sludge, the two samples can be treated separately, as in the case

Jan 1, 1931

By F. T. Havard

The present truce in litigation between Western smelting and ranching interests gives opportunity for a summary of the results achieved by metallurgists in condensing fume and de-acidifying furnace-gases. In the absence of complete records of past experiments, there ha's been a good deal of unliecessary duplication of attempts and failures. Yet, especially uilder the recent policy of our large corporations, looking to the economic recovery of the values of by- and waste products, . original and, in some cases, effective measures have been developed by American metallurgists. In the case of Bliss us. the Anaconda Copper Mining Co., the Montana Suprenle Court, recognizing the progressive efforts made to render the flue-gases innocuous, and the value to the community of the smelting company's operations, refused to hamper its work, and returned a verdict in favor of the defendant. In the trial, the experimental farms of the company ill the so-called "aff'ected region" were proved to be thriving perfectly, showing that immunity was guaranteed to vegetation and stock by tlle wide distribution of the sulyhurous acid gas in the dry air of Montana. Indeed, L. S. Austin has proved that by the time the fume is 4,000 ft, distant from the stack, the percentage of sulphurous acid has been diminished from 2.28 to 0.00045; and Angus Smith has shown that rain must contain 40 parts per million of free acid before vegetation is affected. The arsenic in the stack-gases is but a very small part of the original content of the ore, of which 80 per cent. and over is rejected in the coilcentrator and furnace waste-products or recovered in the flues, Nevertheless, neither the peace in Montana nor the armistice

Jan 1, 1911