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By J. L. Wyatt

THE smelting of ilmenite to produce a slag rich in titanium, with pig-iron as a byproduct, introduces new concepts in electric smelting metallurgy. Titanium slags are characterized by low electrical resistance, thereby requiring a different type of electric furnace operation, and power supply sources substantially different from those encountered in normal steel melting and siliceous ore smelting furnaces. This paper describes special apparatus for determining specific electrical resistivity of titanium slags, and reports the values obtained on slags encountered in the process of smelting ilmenite ores. In an induction furnace under conditions approximating normal furnacing operation, 25 general heats and 6 special standardization and comparison heats were run. Results indicate relatively low resistivity. Resistivity depends primarily on slag composition, and, to a lesser degree, on temperature. During the summer of 1947 a cooperative project between the National Lead Co. and the Bureau of Mines was conducted at Boulder City, Nevada, in the Electrometallurgical Laboratories of the Bureau. The project involved smelting of ilmenite ores in electric furnaces, and has been reported in detail.' The author participated as a member of the staff of the National Lead Co. during the entire campaign. One outstanding characteristic of the titaniferous slags, their apparently excellent and variable electrical conductivity, was considered sufficiently important to warrant further study. Although little electric smelting of iron ores has been done in the United States, a certain amount of information is available for comparative purposes. Personal experience in smelting siliceous iron ores in electric furnaces furnished further information for comparison. In general, silicate slags are relatively poor conductors of electricity. Current flow during smelting operations may be maintained through a floating layer of coke on top of the slag bath, or through a bath of molten iron, beneath the slag. Silicate slags are not sufficiently conductive to furnish a path for current passage when normal furnace voltages are used. Experimental smelting of a silicate iron ore was undertaken for comparative purposes. A fluid melt was not obtained until the FeO content was reduced to 15 pct. The resistivity increased as the FeO content decreased, FeO compounds being the most conductive of the various oxides in such a normal slag system. The converse appears to be true in slags containing large percentages of TiO2. It has been found that titanium slags become molten with FeO percentages in excess of 30 pct, and that the conductivity becomes generally better as the FeO con-tent of the slag diminishes. This phenomenon was demonstrated in the furnacing operations at Boulder City. During the initial smelting period furnace electrodes could be held immersed several inches in the slag, with tap selectors set for a no-load voltage of 70 v. As the FeO content decreased, the furnace electrodes would "walk up," until at one point in the process the tips would break the surface and arc on top of the slag. This was undesirable since direct arcing caused losses of important flux additions. The slag compositions under study at that time were such that losses due to surface arcing sometimes caused viscous slags to be formed. Forcing the electrodes into the slag bath by manual control to prevent surface arcing resulted in excessive power input and lack of control. Suitable low voltages were not available. Measurement of the electrical resistivity of titanium slags at various operating temperatures and slag compositions was undertaken. It was believed that this information would be of use in the design of electric furnaces suitable for smelting ilmenite, with voltages, electrode sizes, and electrode spacing adapted specifically to titanium slags.

Jan 1, 1951

By T. B. Counselman

An iron blast furnace of 1000 tons daily capacity will produce a 100,000 cu. ft. per minute of blast-furnace gas. This contains a 25 per cent of carbon monoxide, and has a B.t.u. value of about 95. gas is used primarily for heating the stoves of the blast furnace, secondarily, under boilers, for steam production. Occasionally it is used in gas engines. The raw gas coming out of the furnace carries about 10 grains of dust per cubic foot of gas, under best operating conditions. With furnace "rolling" or slipping, this quantity may temporarily be gre increased. This dust would, of course, deposit on the checkers o: stoves, or plug the burners under the boilers; therefore the gas : be cleaned. Flow Sheet The first step in the cleaning operation, as shown by the flow s Fig. 1, is a dry dust catcher, of which there are several forms, partially cleaned gas next goes to a primary wet washer, where v sprays knock down most of the solid particles. If further cleani desired, the gas may then go to electrical precipitation. Again, the removed is collected in water. Dry dust catchers reduce the dust col from 10 grains to between 4 and 6 grains per cubic foot, and rec roughly half of the total flue dust in the gas. The wet washers rea most of the remaining dust, reducing the dust content to about 1/4 per cubic foot. Electrical precipitators, if used, further reduce the content to about 0.01 grain per cubic foot of gas. The flow sheet s the dust content of the gas at various steps in the treatment and als content of solids in the water. It should be noted that some plant less water on the primary wet washer than indicated, and that ofte solids content of the scrubber water will be 300 grains per gallon, or much higher. Reasons for Recovering Flue Dust.—There are two main incentive recovering flue dust from the scrubber water: (1) the value of the (2) stream or lake pollution, where such a reason applies.

Jan 1, 1937

By R. L. Curfman

WHEN the Uranium Reduction Co. mill was put on stream in October 1956, one of the many operational problems was that the sand-slime separation circuits could not produce a satisfactory RIP feed, inasmuch as production greatly exceeded de- sign capacity. The original cyclone portion of the circuits consisted of two stages of three 14-in. cyclones in each circuit, as shown in the condensed flowsheet. The high percentage of +325 mesh material in the over- flows of these 14-in. cyclones was not consistent with good operation, and flow rates per ton of ore were high, indicating that not enough of the total solids in the ore was being rejected as sands. It was inferred by members of the metallurgical department that returning the underflows, or sand fraction, from these cyclones back into the drag classifiers was the primary reason for dirty feed.

Jan 7, 1958

By Victor G. Hills

CONTRARY to what seems to be the general impression, the island of Cyprus was not named for the metal copper, but the reverse was the case. The origin of the name is entirely lost. The ancient city Kition, the birthplace of Zeno, is supposed to be the same as Kittim mentioned in Gen. x, 4; Jeremiah ii, 10 and Ezekiel xxvii, 6, and Josephus states that this name seems to have been applied to the city and also to the island itself. Cyprus is 50 miles off the coast of Asia Minor. The mountains on the mainland are plainly visible across the sea from where I write six miles from the shore of Morphou bay. It is 200 miles from Joppa the port of Jerusalem, and 250 from Port Said. The shape of the island is like a lizard, the tail pointing directly toward the old city of Antioch near the northeastern corner of the Mediterranean. The island was once heavily timbered. Among the great trees the cedars were said to even excel the famous cedars of Lebanon. The Odyssey refers to it as the "sweet smelling island" and makes other references. But the continual wars of every race and age completely stripped it of its timber, and for centuries it has been woefully barren. There are two ranges of mountains. One borders the north coast and runs out in a slim rocky ridge forming the "tail." The other occupies the southwestern part of the island, the highest point being Mt. Troodes 6600 ft. high, sometimes called Mt. Olympus. Whether or not it be the Mt. Olympus, anyway right around there the ancient gods and godesses frequently "started things." These peaks are in plain sight 10 miles south from where I live.

Jan 1, 1926

By M. C. Flemings, P. J. Ahearn

Dendvite morphology of unidirectionally solidified Sn-12 pct Bi alloy was determined with the aid of X-ray and macroetch techniques. Columnar growth direction (growth direction of primary arms) in this alloy is [110]; secondary arms grow in the [lii], [111), and [112] directions. Interstices between [112] dendrite arms fill in during early stages of solidification so that well-defined (110) planes are seen in the final structure (parallel to the heat-flozo direction). Similarly, interslices betweex the [111] secondary arrlls fill in to form (112) planes, and interstices between [1111 ] arms fill in to form (112) planes. DENDRITE morphology in alloys is conveniently studied in unidirectionally solidified castings. In this type of solidification, heat is extracted from a single chill surface of the casting while other surfaces are insulated. Solidification progresses from the chill and results in cast grains that are columnar with the long axis of the grain perpendicular to the chill. Dendrite morphology is conveniently studied by examination of etched surfaces parallel and perpendicular to the chill, as previously described.'" Experimental studies on the growth direction for tin are in general agreement. The major growth direction is [110]:-' although this is altered somewhat by supercooling. In one study, tin dendrites pulled from a supercooled melt had a growth orientation 12 deg away from the [110] direction toward the [002] direction.' This orientation could be changed back to the [110] or close to the [110] direction by changing the tempera- ture gradient in the melt. In a study using autoradio-graphic techniques, (111) directions were shown to be directions of preferred growth in addition to the [110) direction.7 PROCEDURE The samples employed in this study were cut from 12 pct Bi-88 pct Sn castings, 1.5 in. in diam and 3.5 in. tall. The casting procedure consisted of heating the charge to 390°C and then pouring the molten alloy directly into a gypsum sleeve mounted on a massive copper chill. Specimens were cut from the ingot, mounted, polished, and etched with Taffs reagent [FeC13 .6H2O, 2 g; HC1, 5 cu cm;H20, 30 cu cm; CzHaOH (95 pct), 60 cu cm].' The dendritic pattern was clearly visible after etching. Photomicrographs at about 8 and 35 times were taken on surfaces parallel and perpendicular to the chill. Laue back-reflection patterns were obtained from three different columnar grains in a singIe polished face 2 in. from the chill. The cut face was polished and positioned in the beam of a Laue camera, so that diffraction occurred from a single dendritic grain. The resulting diffraction pattern was interpreted to determine crystal log raphic orientation in relation to the heat-flow direction and in relation to microstruc-ture of the given grain. RESULTS Fig. 1 shows a horizontal and vertical section of the columnar structure, at low magnification. Light regions are rich in tin, and darker regions tin-poor; the primary phase is tin-rich a phase. The two sections shown were taken orthogonally from the same metallographic specimen. It is readily seen from Fig. 1 that the dendrites have

Jan 1, 1968

By J. Fred, Johnson

THIS TALK is a digest of some of the information contained in tables, one on practices and one on costs of shaft sinking, in Bulletin 357 of the U. S. Bureau of Mines written by E. D. Gardner, Supervising Engineer of the Bureau at Tucson, and myself. The table of practices contains information on 43 shafts and the table of costs contains information on 25 shafts. These various shafts were sunk between 1917 and 1930 so they may be termed modern practice with costs applicable in the United States, Canada and Mexico. In the compilation of this material the statistics on shafts which were concreted have been eliminated, as these figures were entirely out of line with corresponding figures for timbered shafts. It is my experience that where costs are divided between sinking and concreting the tendency is to load the costs onto the larger item of concreting. Thus a shifting of costs making an addition of 10 per cent in concreting costs may suffice to make a deduction of 40 per cent in sinking costs. Fine work is done (on the books) and everybody is happy. Most shafts are sunk with three shifts of men per day. Each shift consists of men who are miners, muckers, or timbermen, as chance may call for during their shift. Where pumping or high fixed costs do not necessitate three shifts it is cheaper to sink on two shifts. It is desirable for speed and economy in any case to work out a cycle of operations for each 24 hours. In this way men are hired for their ability as machinemen, muckers, or timbermen; they know exactly what to do and there is no passing the buck to the next shift.

Jan 1, 1933

By H. E. Wilborg, G. V. Henderson

This chapter deals with those sands employed by foundries for the manufacture of cores and molds used in the casting of such common metals as steel, gray iron, ductile iron, aluminum-based alloys, and copper-based alloys. Most metal is cast in "green sand" molds in which the sand mixture, consisting primarily of silica, clay, and water, forms a plastic material. This plastic material is rammed around a pattern in a split box called a flask, then the box is separated, the pattern is withdrawn, and the metal is poured. These green sand molds are broken apart when the metal has solidified. The sand is then remixed and a new mold made for the next casting. To a lesser extent silica sand is mixed with resins or oils and, after forming, the mixture is hardened by either heat or chemical reaction. In any case, the foundry mold or core is used just once before being broken up and remixed for manufacturing another core or mold. By far the greatest consumption in the foundry is of those types of sand high in silica, low in clay (under 0.5%), and occurring in numerous locations thoughout the United States. Consumption End Uses Foundry sands fall into four distinct categories: silica sand, lake sand, bank sand, and natural molding sand. Silica sand is a general term used to describe washed, graded, dried, and cooled clay-free sands which are found in many parts of the United States. They generally contain 90 to 100% silica but vary greatly in chemical analysis and grain structure. Lake sand is the sand dredged from Lake Michigan or dug from the dunes bordering that lake. Bank sand is composed primarily of silica grain but containing up to 5% clay. It is shipped in both the damp and dry state but the normal processing of bank sand consists only of drying and scalping. Occasionally two different grades of bank sand are blended prior to or during loading. Natural molding sand is a term used to describe a sand containing sufficient clay as mined to be suitable for molding purposes. Clay content may vary from 10 to 24% with the balance of the material consisting primarily of silica sand. These sands range in fineness from one peaked on the 20 mesh to one which contains more than 50% of the particles passing the 270 mesh. Lake sands are relatively coarse and are used in great quantities for the production of automotive castings. Bank sands and natural molding sands are gradually phasing out as the production of castings becomes more a science and less an art. The vast majority of foundry sands used today are those known as the silica sands and lake sands. Foundry sands are used in two basic ways. The majority are used in the green or uncured state with the balance used in the cured or set state. Green sand is grain bonded with clay and contains varying amounts of water to plasticize the clay in order to bond the grains together. Green sand is either naturally bonded or "synthetic" sand. The naturally bonded molding sands are generally found in the Hudson River Valley of New York State, Cumberland County in New Jersey, and in Tennessee and Mississippi. Synthetic sand is molding sand compounded in the individual foundry. It is made up primarily of graded silica sand, western or southern bentonite, water, and carbonaceous and/or cellulosic materials. Green sand is used primarily in making molds. The use of naturally bonded green sand has been declining for many years, primarily because naturally bonded sands are more difficult for the foundryman to find and control. Synthetic green sand, on the other hand, can be

Jan 1, 1983

By M. A. El-Aouar

Drilling muds are complex colloidal systems. In an effort to analyze the physical properties that affect their performance at high temperature, an investigation was made of the hydrothermal stability of the monoionic hydrogen, calcium and magnesium forms of montmorillonite and attapulgite. Samples were heated to 400°1F in a simple autoclave and analyzed for cation exchange capacity, c-axis X-ray diffraction spacing and pH factor us sodium hydroxide titration dependence. Out of the six monoionic clay minerals tested, calcium montmorillonite and magnesium attapulgite showed the least base exchange capacities and were relatively unaffected to heat. It is concluded that, of the six clays studied, these two clays were the most suitable forms for use in drilling fluids. INTRODUCTION The drilling of deep wells is dependent upon the drilling mud, and problems frequently arise concerning the effect of mud temperature and prolonged heating. Drilling muds are complex colloidal systems. In an effort to analyze their behavior in deep, hot boreholes, it was thought desirable in this investigation to study suspension of pure, monoionic montmorillonite and attapulgite clays before and after hydrothermal heating. Chemically, clays are hydrated aluminum silicates, such as montmorillonites and attapulgites. The structure of montmorillonite is composed of units made up of two silica tetrahedral sheets with a central alumina octahedral sheet. Attapulgite consists of double silica chains running parallel to the c-axis with the chains linked together at their longitudinal edges. The apexes of the tetrahedrons in successive chains point in opposite direc-tions.4,7,10 One of the most important characteristics of clay minerals is their property of sorbing certain anions and cations and retaining them in exchangeable state. The property of exchange capacity is measured in terms of milliequivalents (meq) per 100 gm of clay. The cation-exchange capacities are taken at pH 7 from titration curves which show the relation between pH and milliequivalents of added NaoH.9,11 The following are causes of cation exchange: (1) broken bonds around the edges of the silica alumina units would give rise to unsatisfied charges, which would be balanced by adsorbed cations; (2) substitution within the lattice structure of trivalent aluminum for quadrivalent silicon in the tetrahedral sheet and of ions of lower valence; and (3) the hydrogen of exposed hydroxyls may be replaced by a cation which would be exchangeable. The primary techniques employed to study the effects of hydrothermal treatment were titration curves and X-ray diffraction analyses. Titration curves were useful in characterizing the nature of the clay mineral surface and the stability of the various samples to attack by acid or alkali before and after hydrothermal treatment. X-ray diffraction analyses were used to determine changes in the nature and crystal structure of the minerals present and to determine the cations in exchange positions on the montmorillonite samples. EXPERMENTAL PROCEDURE PREPARATION OF SAMPLES Selected samples of Wyoming bentonite and attapulgite clay from mines along the Georgia and Florida border were purified by centrifugation to remove all particles larger than 2 microns. In preparing magnesium attapulgite, first a 1.00-normal solution of magnesium acetate was prepared. The pH of the solution was adjusted to 7.0 by adding magnesium oxide. One hundred grams of attapulgite clay were mixed in a blender for 30 minutes with the magnesium acetate solution. The attapulgite solution was left overnight, and then filtered in a Buchner funnel. The suspension was leached with further quantity of magnesium acetate solution and then washed with distilled water for 20 days, after which the filtrate was free from magnesium ions. The concentration of magnesium attapulgite in the clay slurry was adjusted and determined quantita-

By Orson Shepard

THE leaching method that was first widely used with the cyanide process consisted of percolation leaching of crushed ore in vats or leaching tanks. It was frequently necessary to separate the sand for cyanide treatment and discard the slimes, because only granular material allows the necessary amount of solution to percolate through it in a reasonable time. Later a slime process consisting of agitation followed by thick-ening and filtration was developed, and is extensively used to the pres-ent day. Even though the agitation-leaching slime process is considered satis-factory, the cost of plant and of operation are both greater than for a percolation-leaching plant of equal capacity. This situation has been recognized in the hydrometallurgical treatment of copper-bearing slimes; and a considerable amount of j work has been done at the Southwest Experiment Station of the U. S. Bureau of Mines in an effort to develop a percolation method of treating slimes. Experimental work on the copper-leaching problem was started by H. E. Keyes1, and later tests were made by J. D. Sullivan and A. P. Towne2. The Bureau of Mines work resulted in the development of a process by which slimes were moisture-agglomerated with sand and crushed ore, to form a permeable mass through which solution could be trickled without breaking clown the glomerules. The agglomeration trickle-leaching method was devel-oped on a small scale and finally tried out in a 12-ton vat. Successful results were obtained as long is the leach solution was added slowly, so that the agglomerated charge was never flooded. If the leach solu-tion flooded the charge, the glomerules broke down and, the charge became impervious. There are two important difficulties in the application of the agglom-eration trickle-leaching-method to the cyanide process for gold and silver ores:

Jan 1, 1937

By S. F. Frederick, W. A. Bailey

The relationship between microstructure and mechanical properties of cast 356-type aluminum alloys was studied to determine the cause of the variations in properties resulting from differences in solidification rate. It was found that variations in strength are a consequence of variations in ductility and that ductility is inversely proportional to the dendrite cell size. A mechanism is proposed to account for this correlation based on the fracture strain of the inter-dendritic silicon particles and the differential strain across dendrite cell boundaries. IT has long been assumed that permanent mold castings of aluminum are superior to sand castings because of their finer grain structure and reduced porosity.' A systematic variation of static properties was noted during unpublished studies of fatigue-strength variations with solidification rate of A356-T6. Of particular interest was the observation that for a given composition and heat-treatment, both ductility and the ultimate strength increased with higher solidification rate. Because the A1-Si system forms the basis of many alloys used for castings and welding wire, a study was initiated to determine the details of how higher cooling rates improved mechanical properties of these alloys. The results of this study are presented here. EXPERIMENTAL PROCEDURE Specimens were prepared from 356-T6 to supplement the A365-T6 alloy data from earlier studies. This allowed a comparison of behavior of alloys with substantially different strengths but with only small differences in alloy composition. The composition and heat treatment of both alloys are given in Table I. The specimens were cut from a chill plate casting,' which is a 6 by 9-in. plate, 1 in. thick, which has one end heavily chilled, and a large riser at the other so that solidification is essentially unidirectional from the chill end to the riser. Although not measured here, solidification times measured on similar specimens by embedded thermocouples are a few seconds at the chill end and up to 5 min near the riser. Specimens were obtained by cutting slabs from the plate, parallel to the chill; these represented different cooling rates, with essentially the same composition. For the tests, either 0.125 or 0.250-in. thick slabs were cut for standard 2-in. gage length specimens. The thicker specimens were tested to failure at 2000 lb per min while the thinner specimens were tested in an Instron machine at a constant cross-head speed of 0.02 in. per min. These thin specimens were intended for trans- mission electron microscopy. Some thin specimens were tested to failure while the remainder were given various amounts of prestrain. Dendrite cell sizes were measured by the linear intercept method on photomicrographs at either X100 or X50, depending on the cell size. It was assumed that the dendrite cells were outlined by the silicon particles. Determinations were made on two mutually perpendicular planes and an rms average taken. Electron fractographs of representative tensile specimens were made in the conventional manner using the two-stage replica technique.3 Samples for transmission electron microscopy were prepared by chemical milling followed by electropolishing. RESULTS AND DISCUSSION The results of the tensile tests and the dendrite cell-size determinations are given in Figs. 1 and 2. All data show the same trends, namely, a decrease in ultimate strength and elongation with increasing distance from the chill (and solidification time), and a corresponding increase in dendrite cell size. The yield strength remains essentially constant. The constancy of the yield strength was particularly significant since it indicated that variation in the solidification rate does not affect the stress required to initiate plastic flow. By plotting true stress-true strain curves as in Fig. 3, it can be seen that both

Jan 1, 1969

By G. J. Ballmer

Drilling and blasting 12-in. blastholes started about the middle of 1946 and has worked out so well that about one half of the blasting, formerly done with 9-in. holes, is now done with 12-in. holes. In most of this work the harder benches were chosen for the 12-in. size because these were the areas where the greatest difficulty was encountered in thorough fragmentation. At the start of this experimental work a 42-T Bucyrus- Erie churn drill was available whereas all other blasthole drilling was being done with 29-T rigs. The large drill was first equipped with a standard 9-in. string of tools but it could not drill quite as fast as the 29-T drills. How- ever, when the weight of the string was increased about 700 lb, or to a total weight of about 3500 lb, it was found that the 42-T exceeded the drilling speed of the 29-T by about 25 per cent. When a 12-in. bit was substituted for the 9-in. bit the 42-T rig again dropped behind the drill- ing rate of the 29-T rigs. At this stage of the experiment the 42-T was using a 12-in. bit with detachable points and the total weight of the string of tools was about 3800 lb. It took longer to change the drill points than it did the regulation bit and this accounted for some of the slower drilling rate. When a heavier stem and standard construction 12-in. bits were used it was found that the 42-T could drill a 12-in. hole slightly faster than a 29-T could drill a 9-in. hole even in the hardest rock in the pit.

Jan 1, 1949

By L. Moyd, P. Moyd

The first commercially available portable gamma ray-neutron beryllium detector, the Rerylometer, was developed by the same group that developed the first practical portable scintillation counter, the Scintillo-meter. This group includes Professor George Brownell, of the Geology Department of the University of Manitoba in winnipeg, and his associates, K. I. Roulston, R. w. Pringle, and R. L. Funt, of the Physics and Engineering Departments of the University of Manitoba. A prototype model was tested in the field by Dr. Brownell during the summer of 1958. Commercial production was undertaken shortly thereafter by Nuclear Enterprises Ltd., of Winnipeg, and a license for manufacture and distribution in the U.S. was granted to Isotopes Specialties Co., of Burbank, California, a division of Nuclear Corporation of America. Early in 1959, we were retained by the 'Mining Division of Minerva Oil Co. to Legin a search throughout North America for minable deposits of beryllium, and to train members of their permanent staff to continue the program. The project included literature sur-veys, interviews with members of various federal, state, and provincial agencies, scanning of our own collection of suites of specimens from various mineral occurrences, and the examination of deposits in the field. The value of a beryllium detector to this program was obvious and arrangements were made to procure one as soon as possible, which resulted in our obtaining and using the first instrument that was produced commercially in the U. S. In addition to the commercially available Berylometer, Mr. William Vaughan and his associates, of the U. S. Geological Survey Geophysics Division, in Denver, have developed and tested several experimental instruments based on the same principles. DETECTOR Principles The detection of beryllium is based on the fact that "hard" gamma rays (1.6 MEV and geater), such as those produced by Antimony convert ordinary beryllium (Be9) to Be8 , with the release of a neutron. The neutrons are counted by a scintillator designed to favor neutron responses while suppressing those of gamma rays. 'The instrument is completely selective for beryllium. No other element can interfere. Instrunlentation The Rerylometer itself, which is cylindrical, is about 8 inches in diameter, 13 inches high, and weighs about 11 pounds. (Figs. 1, 2). The principal components include: 1. A disc of wax to moderate (or "thermalize") neutrons. 2. A boron-enriched polyester-zinc-sulphide fluorescent screen. This material produces strong flashes when struck by thermalized neutrons but only weak flashes from gamma rays. 3. A photomultiplier tube. 4. An electronic circuit, complete with one srnall 71/2-volt dry battery. The photomultiplier tube and electronic circuit convert neutron flashes to electrical impulses, but gamma ray effects are almost

Jan 1, 1961

By J. W. Eckerd, P. S. Lewis, N. H. Coates

USBM designed, constructed, and operated an 18-in.-diam fluidized-bed combustor for highly caking coals to evaluate the method for possible application to power generation. In initial tests, combustion was carried on successfully, with carbon losses as low as 0.8%+ Coal ash was not satisfactory as a bed material, but mullite proved to be acceptable. Overall heat-transfer coeffients from the bed (-20 t 50 mesh mullite) to a water-cooled tube were about 75 Btuper hr-sq ft. Fluidized-bed combustion is being investigated by the U. S. Bureau of Mines (USBM) as a potentially cheaper method of firing power plant boilers. Lower combustion temperatures in fluidized beds—only 1400" to 1800°F-—should reduce fireside deposition and corrosion and nitrogen oxides production. Fluidized-bed combustion also offers possible advantages in the use of additives to control SO, emissions and amenability to processing of high-ash, high-moisture coals without special preparation. This paper describes development of a fluidized-bed combustor and the results of preliminary tests. Equipment and Procedure Fig. 1 shows the 8-ft-high combustor. A cone-shaped plate perforated by I/8-in.-diam holes supports the fuel bed and admits the fluidizing air; water passing through the 3/4-in. pipe extracts heat from the combustion bed. Fig. 2 is a flow diagram of the system. Coal is fed pneumatically to the base of the fuel bed. Combustion products are passed through two centrifugal separators to remove most of the entrained solids, then to a water

Jan 1, 1971

By Frank C. Smith

THE following article covers the history of a metallurgical campaign, commenced in March, 1905, at the mines of the Socorro Gold Co., in the so-called desert region of Yuma county, Arizona. The results obtained were simply a function of necessity, as will be seen from the description of the local economic conditions and the character of the ores treated. The mines and mill of the Socorro Gold Co. are located near the small village of Harrisburg, about 60 miles from Congress Junction, Arizona, which, at the time this work was undertaken, was the nearest railroad station, and from which all supplies were brought by wagon through the desert, at a cost varying from $10 to $15 per ton in addition to the railroad freights. Wood, the only available fuel, cost $7 per cord. Water for all purposes was pumped to the plant from a well in the valley about 4 miles away. Under the Arizona law, a working-shift is 8 hr. Wages were $6 per day for foremen and mill-men, $4 per day for blacksmiths and engineers, $3.50 per day for miners, and $3 per day for shovelers and trammers. The ore consisted of quartz, in which, above the 250-ft. level, the iron-minerals were largely oxidized and some free gold was visible; below that level few traces of oxidation occurred, and pyrite constituted the principal mineralizer in the quartz, together with occasional pockets of galena and a few eccentric specks of covellite. The 20-stamp mill was equipped with plates for amalgamation, and three Standard tables for concentration. The tails from the tables were elevated by a centrifugal pump to a launder, by which they were conveyed to a 60-ton cyanide-mill for further treatment. The cyanide-mill was erected at so great a distance from the stamp-mill that a separate crew of workers and a separate power-plant were necessary. No arrangement had been made for the treatment of the concen-

Jan 1, 1907

By A. E. Lillstrom

IN the development of iron mines and production of iron ore from the Marquette range, drilling blast-holes is an important phase of the mining cycle. The ground drilled in ore production can be classified into two main categories, soft hematite and hard hematite or magnetite. Within these categories the material exhibits a wide range of penetrability by percussion drills. Development work encounters various types of rock. Slate and altered basic intrusives constitute the softer types commonly encountered. Harder materials are represented mainly by greywacke, quartzite, iron formation, and diorite. Prior to the first tungsten carbide trials in late 1947 and early 1948, hard-rock and ore drilling was done with steel jackbits starting at 21/4-in. diam. These were reconditioned by hot milling. Automatic or handcrank 31/2-in. drifters were employed, mounted on Jumbos, posts and arms, or tripods, depending upon the working place. With the exception of shaft sinking jobs where 55-lb sinker machines were and still are used with 1-in. quarter octagon steel, the other production and development mining utilized 11/4-in. round and Leyner-lugged steel. The following properties have been selected as typical examples wherein carbide bit applications have proved economical. The Mather mine "A" and "B" shafts and Cleveland-Cliffs Iron Co. mines are soft ore mines where insert bits are used in rock development only. The Greenwood mine, Inland Steel Co., Champion mine, North Range Mining Co., and Cliffs shaft mine, Cleveland-Cliffs Iron Co., are hard ore mines where all drilling is done with tungsten carbide bits. Mother Mine "A" Shaft In the Mather mine "A" shaft and other soft ore properties where only rock development work is done with the tungsten carbide bits, several types and makes of bits have been tried since early 1948. The greatest proportion of failures have been at the connection end, although the early trials with the 13 Series Carset 11/2-in. bit used in conjunction with 31/2 -in. automatic-feed drifters, showed an equal amount of shattered inserts. To combat this shattering, the 31/2 -in. drifters were replaced by 3-in. drifters, thus eliminating, for the most part, insert failures. However, the attachment end of the rod continued to be the main source of trouble. The greatest amount of failure was in the stud or at the upset section approximately 2 in. behind the drive shoulder of the rod. Heat treatment was changed several times as well as the composition of the alloy studs. Since this failed to correct the trouble, a decision was made to change to a heavier attachment section. Timken 11/2-in., type M, bits were then employed and showed an exceptional improvement. The rods are discarded when the thread contour shows sharpening or wear on the shoulder. It was also learned that the Timken insert did not show as rapid gage and cutting edge wear as did competitive makes, and footage per use increased by approximately 50 pct. Prior to the Timken trials the average life per bit at the Mather mine "A" shaft on 6-ft change chain-feed drifters was 500 ft, and the rod life at the connection end was 50 ft. The Timken bit with chrome-plated thread averaged 1200 ft, and rod life increased to as much as 500 ft. However, the life of the connection end was much better on shorter length drill rods or in places where machines with 34-in. change were used. The bit thread continued to be the point of ultimate failure with thread strippage, constituting the cause for discard of bits. In one of the new development headings, harder rock was encountered for approximately 800 ft, dropping the life per bit to a low of 90 ft with shank and thread life of rods dropping to approximately 125 ft average. The stripped bits were then welded to the rods, increasing the life per bit by 75 to 100 pct. The rod transportation for main level development was not a problem so intraset rods were tried. Intraset rods have tungsten carbide inserts set into the rods proper by the manufacturer and can be obtained with chisel or four point bits. This type of rod eliminates the need for any connection and the steel being a special alloy will show more feet drilled per rod. The first trial was made with eight rods, and final results averaged 350 ft per rod, six of the rods worked the life of the bit end, and two broke shanks at less than 50 ft. The preceding example showed a considerable improvement, so additional steel of the same type was purchased, but its use has been limited to main level drifting only, because of the handling problem involved in transportation of the complete rod to mine shops for resharpening. Further trials are being made on improving the life per detachable bit by chrome plating. To date, the chrome plating shows an improvement of approximately 100 pct. However, final results will not be known until the present long term trials have been completed. Mother Mine "B" Shaft In November 1947, tungsten carbide bits were first tried at the Mather mine "B" shaft. The use of 1%-in. Carset 13 Series bits, for drilling the 72-hole, 7-ft shaft round, decreased the drilling time from an average of 41/2 hr per round required with steel bits, to 2 hr with insert bits. The best drilling time for

Jan 1, 1952

By P. H. Royster

IN a study of the blast-furnace process, the Bureau of Mines has made many experiments for the purpose of determining the exact nature of the combustion of coke in the neighborhood of the tuyeres. Two papers1 presented at a meeting of this Institute discussed somewhat the practical aspects of coke combustibility in its relation to the operation of the blast furnace. In a later paper,2 Sherman and Kinney discussed the possible effect of variations in the combustibility of cokes on the furnace operation, and find that the "reoxidation theory" of Koppers and the "heat concentration theory" of Brassert3 are insufficient to explain the variation in the conduct of different cokes that has been observed by furnace operators. In their report to the D-5 Committee of the American Society for Testing Materials, with which the Bureau of Mines has been cooperating, Perrott and Fieldner4 seem to agree with Sherman and Kinney that, for practical purposes, the nature of the combustion of coke at the tuyeres of any furnace is constant and is affected neither by the physical properties of the coke nor by the nature of the furnace operation. They omit as unimportant any tests on the combustibility of coke from their proposed methods for standard tests on coke. In the last two papers, it is pointed out that performance tests, in which all other factors except coke quality are kept constant, are to be desired and

Jan 3, 1924

By R. A. Garling

INTRODUCTION Since the infancy of in situ uranium mining, a growing number of hydrometallurgical processes have been incorporated into pilot and commercial scale flowsheets. Although initial design efforts were geared toward maximizing uranium recovery and minimizing plant and wellfield flow circuit maintenance, recent emphasis has shifted to improved means of water conservation and aquifer restoration. As mining units approached depletion, evaporation ponds reached minimum freeboard, and state and federal agencies demanded proof of groundwater restoration, processes including mixed bed and conventional ion exchange, reverse osmosis and electrodialysis were adopted by the industry. These units served the additional function of reducing process bleed flows during mining in states where the deep disposal well permitting ice remains unbroken. This report concerns the use of electrodialysis as an alternative to the more conventional processes used in in situ mining. In addition to a brief history and description of the process, a comparison to reverse osmosis and operational data derived from testing an Ionics, Inc. 1.31 x 10-3 m /s (30,000 gallon/day) unit at the Teton-Nedco Leuenberger Research and Development pilot will be presented. HISTORY Commercially practicable electrodialysis was contingent upon the development of synthetic ion exchange membranes in 1940's. In 1952, Ionics Inc. demonstrated that the process was amenable to the treatment of salt and brackish water and, in 1954, made their first commercial sale. The following decade saw several major electrodialysis unit sales which were generally targeted for use on private or municipal potable water treatment. Major increases in membrane desalting unit capacities, facilitated by technological advances in the reserve osmosis industry, were noted during the 1970's. The development of polarity reversing electrodialysis equipment which reduced feed pretreatment requirements, increased water recovery rates, and simplified unit operation, kept Ionics Inc. competetive in the water treatment industry. Engineering advances which incorporated automated equipment, non-corrosive construction materials, and improved ion exchange membranes allowed the electrodialysis process to compete in industrial waste treatment among other commercial markets. PROCESS AND APPARATUS DESCRIPTION The electrodialysis process utilizes direct electrical current passed across a stack of alternating cation and anion selective membranes in order to achieve an electrochemical separation of ionized materials in an aqueous solution. The membrane stack has the appearance of a plate and frame filter press and auxilliary equipment includes solution pumps, electrically actuated valves, filters, piping and a direct current power source. The ion separation membranes are thin sheets of synthetic cation or anion selective resins. Attaching sulfonate or quaternary ammonium groups to the cross linked copolymer structure determines the ion selectivity of the membrane. The membranes are separated from each other in the stack by non-conductive spacers that house flow channels which route the flow tortuously and parallel to the membranes. Direct electrical current passing perpendicularly to the membranes and solution passages attracts cations toward the cathode and anions toward the anode (Figure 1). As the ions from the feed stream pass through the ion selective membranes, they become concentrated in the adjacent brine channel and are retained there by the combined attractive force of the electrode and the repelling force of the next membrane toward the electrode. Limiting factors on the degree of demineralization possible include chemical solubilities in the brine flow and the current density that will produce an unacceptable degree of polarization (Figure 1). Feed or brine solution treatment with complexing agents or acids has been successfully applied to prevent membrane scaling. Polarization can occur when sufficient current density is applied to dissociate water in the ion depleted region of the diluting compartments near the membrane surfaces. Significant polarization is evidenced by large electrical resistances across cell pairs and notable pH differences between diluting and concentrating streams. Limiting current densities have been increased in U.S. manufactured equipment by utilizing tortuous flow paths of relatively high linear velocities thereby promoting continous solution mixing. Energy consumption is due to separating electrolytes and solutions, oxidation and reduction reactions occurring in electrode compartments, overcoming electrical resistance, conversion from AC to DC power, solution pumping and auxiliary equipment actuation. A major improvement to the basic electrodialysis process was applied in 1970 which resulted in frequent, automatic cleaning and descaling of membrane surfaces. The process, polarity reversal, incorporates alternating the cathode and anode on a periodic basis while exchanging product and brine flow channels via electrically actuated values. The reversal reduces the potential of stack plugging with CaCO3 (calcite), CaSO4 (gypsum), and colloidal materials and, in most waters, eliminates feed pre-treatment requirements. For approximately two minutes during and following the reversal, off spec. water is flushed to waste or reintroduced to the feed supply. The usual feed treatment on polarity reversing electro-

Jan 1, 1982

By P. H. Royster, T. L. Joseph

In a study of the blast-furnace process, the Bureau of Mines has made many experiments for the purpose of determining the exact nature of the combustion of coke in the neighborhood of the tuyeres. Two papers1 presented at a meeting of this Institute discussed somewhat the practical aspects of coke combustibility in its relation to the operation of the blast furnace. In a later paper,2 Sherman and Kinney discussed the possible effect of variations in the combustibility of cokes on the furnace operation, and find that the "reoxidation theory" of Koppers and the "heat concentration theory" of Brassert³ are insuffcient to explain the variation in the conduct of different cokes that has been observed by furnace operators. In their report to the D-5 Committee of the American Society for Testing Materials, with which the Bureau of Mines has been cooperating, Perrott and Fieldner4 seem to agree with Sherman and Kinney that, for practical purposes, the nature of the combustion of coke at the tuyeres of any furnace is constant and is affected neither by the physical properties of the coke nor by the nature of the furnace operation. They omit as unimportant any tests on the combustibility of coke from their proposed methods for standard tests on coke. In the last two papers, it is pointed out that performance tests, in which all other factors except coke quality are kept constant, are to be desired and

Jan 1, 1924

By Thomas T. Read

REVIEWING the events of the year in mineral industry education, a certain amount of either amusement or irritation, depending upon one's viewpoint, can be derived front the section dealing with mining engineering in Walter B. Pitkin's "New Careers for Youth" which appeared during the summer. On p. 102 Prof. Pitkin says "The mining engineer is disappearing .. . because . . . the profits of mining decline rapidly with the age of the mine." In his "Are Too Many Students Taking Mining Courses?" published in the May issue of this journal, W. B. Plank presents statistics showing that there was some decline in student enrollment in mining and metallurgical courses in 1932-3 and 1933-4, compared with 1930-31, but that the decline was proportionate to that in all lines of engineering study combined. Informal reports from Institute officials that have visited tinning schools during the fall indicate that nearly all schools showed an increase in enrollment for the current school year. Pre else statistics will be made available by Prof. Plank at the Annual Meeting. His study last year provided, for the first time, accurate and complete information on the enrollment in mining schools, as well as the number of various courses and the degree granted, and it will be interesting to have comparable figures for this year. Prof. Pitkin's views on the future opportunities for mining engineers were discussed at some length - on pages 439-4.40 of the November issue of MINING AND METALLURGY.

Jan 1, 1935

By O. E., Kiessling

THE better control of production was made the topic for a special program of the annual meeting of the Institute last February. In the discussion at that meeting it was brought out that in many branches of the mineral industry (as in other industries), there was faulty industrial coordination. Those who were best informed by reason of their intimate connection with coal, copper, oil, iron, lead and zinc, and almost all other of the essential minerals pointed to a condition of chronic over-capacity, to the dissipating use of irreplaceable minerals, to an unstabilized price structure, and to over-stimulatecl production and the destruction of some mineral resources over and above what is thought to be the needs of prudent cur- rent consumption. The discussion went far enough to point out the existence of various problems, and the question then raised was "what can be done by way of self-coordination of industry?" by "production control from within?" or in other words, "how can the mineral industry protect its resources against waste and at the same time protect both the producer and consumer against loss from prices that are too low or too high?"

Jan 1, 1928