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By Kenneth Robertson

THERE is probably less known of this subject than of any other connected with the metallurgy of iron. In all the books that treat of this matter, there ore given analyses of slags, their chemical formulae, the loss of iron, etc. ; but few, if any, give the character of the iron mode at the time; none give analyses and physical characteristics of the ores and limestones used, and they present merely isolated examples from various furnaces, and consequently one is more frequently puzzled than enlightened after reading them. During two years I have had occasion to analyze the slags and many of the pig-irons from two furnaces, and for this article I hove made analyses of all the different grades of pig mode by them. A

Jan 1, 1873

By V. F. Swanson

A petrographic study was made of the type and sources of mineral contamination occurring in Florida phosphate rock products. Emphasis was placed on contaminants in the screened rock (+1 mm), since this portion of the product undergoes no physical beneficiation except for scrubbing to remove slimes. Since iron and aluminum contents of the finished product affect the salability, special emphasis was placed on identifying types and sources of these contaminants and relating chemical composition to visual characteristics. Wavellite derived from the overburden was the major aluminum contaminant with some aluminum being included as secondary clay balls. Iron occurred as phosphates in the overburden and as colloidal oxides included in the phosphate pebbles. Chert and limestone contamination derived from the Hawthorne formation underlying the phosphate matrix.

Jan 1, 1972

By E. Gybbon Spilsbury

The recovery of gold and silver from their ores by means of the cyanide process has been so successful in the last few years that ally radical improvement would seem impossible; yet the appliance to which I wish to call attention in this paper is really a radical departure from the methods now in general use. The most modern and approved of these, known as the all-slimillg method, depends for its success on the grinding of the ore so fine that practically 90 per cent. of it will pass through a 200-mesh screen. The slimes thus produced are then agitated and aerated in tanks of various types. The main object of this treatment is to insure such a thor-

Jan 1, 1911

By Y. Wang, M. M. Singh, H. L. Hartman

Knowledge of the state of stress at the bottom of a drillhole is of practical importance in rock drilling, since the induced stresses must exceed the strength of the rock in order to cause fracture. This note reports on a numerical method to determine the stress distribution at the bottom of a borehole induced by a static bit load and overburden and fluid pressures. The procedure adopted is similar to that for the plane-strain, discrete model employed by Harper and Ang.1,2 The rock medium is considered to be homogeneous, isotropic, and elastic-perfectly-plastic. METHOD OF ANALYSIS AND PROGRAMMING The discrete model adopted consists of two types of points: mass points and stress points. The configuration and numbering notation are depicted in

Jan 1, 1967

By Parke A. Dickey

POTENTIAL differences between strata of shale and sandstone have been recognized for about 15 years, and they form the basis of the electrical logging of oil wells. Hitherto these potentials have been ascribed solely to electrochemical reactions caused by the disturbing effects of the water in the drill hole through which they are generally measured. Observations by the author in wells empty of water and in a mine shaft suggest that potential differences between sandstones and shales are natural and specific properties of the rocks, although they are modified to measurement by the disturbing effects of the water in the drill hole. The cause of these potentials probably is to be sought in the relative polar adsorptive capacities of quartz and clay particles.

Jan 1, 1944

By Strathmore R. B. Cooke, Iwao Iwasaki, C. S. Chang

PYRRHOTITE has long been considered a gangue mineral to be eliminated as tailing in the treatment of various sulphide ores. However, in recent years the world-wide lack of sulphur resources has called attention to this mineral as a potential source of both sulphur and iron. Its importance as an economic mineral, however, has not been particularly emphasized. For this reason very little is known about its response to flotation, except that it can be depressed easily in alkaline circuit, by long aeration,1,2 addition of oxidizing agents3 or by starch.4 The object of this work was to study the floatability of pyrrhotite. This includes the effect of oxidation by aeration, of copper activation, and of change in pH.

Jan 2, 1954

MISCELLANEOUS U.S. 0,029,598 - Extraction of vanadium values from vanadiferous iron oxide ore. A pelletized mixture of ore, limestone and bentonite is roasted in an oxidizing atmosphere to form calcium vanadates, the calcine ground, the ground material leached with aqueous ammonium carbonate or sodium carbonate to produce a vanadium-rich solution while leaving an insoluble residue of calcium carbonate and/or calcium bicarbonate along with undissolved iron values, and the solution separated and treated for its vanadium content. This patent is a reissue of U.S. Patent No. 3,853,982, dated Dec. 10, 1974. C. B. Bare and J. W. Pasquali, assigned to Bethlehem Steel Corp. Mar. 28, 1978. 7 pp. (423-68). Filed: 3-3 1-77 (783,352).

Jan 1, 1979

APPROXIMATELY 90 per cent of the steel that is melted and refined in the United States and poured into ingots is made in basic open-hearth furnaces, as shown in Table 1-1. The annual ingot capacity of the steel industry reported January 1,1950, by the American Iron and Steel Institute8* was 99,040,160 tons, of which the basic open-hearth process accounted for 85,896,580 tons, 86.7 per cent of the total. The capacity of the individual furnaces varies over a wide range, a large majority falling within limits of 50 to 200 tons (Fig 1-1). Most of the units constructed [ ] since 1942 are in the larger size groups, between 175 and 225 tons capacity.

Jan 1, 1951

Occasionally it is interesting, and sometimes useful, to review the past for early references to our industry, and to learn of the trials and travail passed through before it arrived where it now is- and the results of such a review follow. Some time ago one of our modern playwrights and poets, Arthur Guiterman, turned to coal, with the following result: Afar in the Carboniferous time I grew in Paleozoical slime, A lepidodendron with boughs on my stem And lepidostrobuses grew out of them. My roots in the mud of the ages I spread And up to the heavens I vaunted my head, I throve in the heat of tropical sun, And now I am coal at two dollars a ton

Jan 1, 1939

By William Price

DURING the past three or four years, the fabricators of silicon-bronze alloys have endeavored to induce risers of phosphor bronze to use instead the silicon-bronze alloys, claiming that the silicon bronzes are superior to the phosphor bronzes for switches and springs, or where high fatigue resistance is desired. This paper gives the results of tests comparing several phosphor bronzes with a 3 per cent silicon bronze and a copper-nickel-tin alloy (Adnic), all processed by the Scovill Manufacturing Co. So far as the authors are aware, sheet-metal fatigue tests have not been made on any of the alloys included in this report, except on one phosphor bronze, material C, Table 1, which was tested by Townsend and Greenall1.

Jan 1, 1937

By C. C. Griggs

FROM its inception, a limestone quarry or mine should be under the direction of a capable engineer. Before it becomes a reality, he should outline the future results, plan the most economical methods for moving the materials and the most efficient equipment, and pilot the undertaking along the narrow pathway of success, avoiding the pitfalls of uneconomical production and the precipitous walls of scanty profits. Inspection of successful limestone properties shows that all have had difficult engineering problems, the solution of which have required keen engineering ability. In approaching a subject of this nature, it is logical to follow the progress of the limestone operation from its inception to the end, as reflecting the engineering required. There are four major divisions: Investigation, equipment, development, and operation.

Jan 2, 1925

Occasionally it is interesting, and sometimes useful, to review the past for early references to our industry, and to learn of the trials and travail passed through before it arrived where it now is - and the results of such a review follow. Some time ago one of our modern playwrights and poets, Arthur Guiterman, turned to coal, with the following result : Afar in the Carboniferous time I grew in Paleozoical slime, A lepidodendron with boughs on my stem And lepidostrobusea grew out of them. My roots in the mud of the ages I spread And up to the heavens I vaunted my head, I throve in the heat of tropical sun, And now I am coal at two dollars a ton.

Jan 1, 1935

By E. A. Lucero, H. E. Robbins, T. H. Dudley

EL Salvador mine of Andes Copper Mining Co., E subsidiary of Anaconda Co., is in northern Chile about 18 miles north of Potrerillos, 7825 ft above sea level. Initial production is planned for the first half of 1959. Plant capacity will be 30,000 tpd, and there are 308,835,000 tons of proven ore reserves, averaging 1.49 pct Cu. The Inca adit is being driven as part of the mine-to-mill haulageway. Because of the overburden it is planned to mine the deposit by block caving. Ore will be dropped from upper gathering levels through ore passes to the Inca adit, through which it will be transported out of the mine to the primary crushing plant located a half mile from the portal.

Jan 3, 1958

In introducing last year's Industrial Minerals Review, Stan Lefond accurately predicted that the 1974- 75 period would be quite difficult and frustrating. Although 1974 demand and prices for most industrial minerals were exceptionally strong, operating costs also rose rapidly-while shortages of supplies and equipment, symptomatic of an overheated economy, occurred from coast to coast. Thus, at the end of 1974 recessionary forces were accelerating-and the minerals industry will soon be facing falling demand as a consequence of sharp declines in auto production, housing, appliances, and other consumer goods. Moreover, when sales weaken, the high-volume, low-prices industrial minerals will incur sharply increased unit costs. To maintain adequate profit margins under these conditions will require cautious and perceptive management.

Jan 2, 1975

By John D. Wiebmer

Lignite Development in North Dakota is a "shotgun wedding" according to former state senator Robert L. Stroup-the unwilling groom (North Dakota) is being led to the altar by the nation's demand for increased coal production. The state's attorney general, Allen I. Olson, reacts to the situation this way: "We have no specific coal policy. We plan to keep the whole matter down to a dull roar and drag our feet until we see how we fit into the national energy picture." But despite a state administration labeled as "cautiously negative" toward resource development, North Dakota has experienced a twofold increase in annual lignite production within the last two years. And within the next four years, three more mine additions will double the state's output again.

Jan 8, 1977

By John D. Hanks

INTRODUCTION Chevron's Uranium Mill is located near Panna Maria, Texas; 70 miles southeast of San Antonio. Designed by Kaiser Engineering, the Mill will process a nominal 2500 dry T.P.D. of uranium bearing ore containing 15% uncombined moisture. Earl Torgerson, San Mateo, California, is the consulting metallurgist on process design. Feed to the plant consists of a mixture of high, medium and low grade sandy day ore; the average grade of the ore will be 0.7% throughout the life of the project. The ore is delivered to the mill via truck and stored by type in individual piles on a flat storage area. The ore is fed by conveyor to a semi-autogenous grinding mill. The SAG mill discharge slurry is pumped either to a storage tank or to the first of five mechanically agitated leach tanks where both H2SO and NaC103 are added. Following leaching, the slurry is mixed with thickener No. 2 overflow before being pumped to a six-thickener, countercurrent decantation circuit where the solution containing uranium is separated from the leach residue. The residue is washed essentially free of solubilized U308 values at the sixth thickener and discharged to an adjacent tailings pond. The first thickener overflow, containing approximately 0.4 grams U308 per liter, is filtered for clarification and sent to the liquid ion exchange (solvent extraction) section. The pregnant aqueous solution is mixed with organic solvent containing amine on which the complex uranyl sulfate ions are absorbed. The immiscible aqueous and organic solutions are mixed and separated in each of the four stages of solvent extraction. The final pregnant organic is directed to a stripping section. A strip solution containing (NH4)2S04 (ammonium sulfate) and NH4C1 (ammonium chloride) is contacted and separated from the organic in each of the four mixer-settler units. The strip solution is mixed with NH3 and the uranium precipitates along with trace amounts of sulfate, chloride, and ammonia. After washing in a thickener, the uranium precipitate or yellowcake is centrifuged and dried in a multiple hearth roaster. Overflow from the yellowcake thickener is recycled back to the stripping section of the solvent extraction circuit. The dried yellowcake concentrate contains more than 98% U308; diluents include H20, ammonia, chloride, etc.. Impurity concentratons in the product are sufficiently low, after drying, to permit direct shipment to refinement installations. ORE RECEIVING Ore from one or more mines will be stored on a pad adjacent to the uranium processing mill. Ore stored in the area may total 200,000 tons or more, which is equivalent to over a two month treatment reserve for the mill. In addition to providing surge capacity, the proposed storage facility permits natural oxidation of the ore and affords an area which, in turn, improve plant U308 recovery and reduces consumption of oxidizing reagents. The uranium bearing ore is segregated at the storage area into several distinct types. One group can be identified by its sandy, clay-like matrix and will be distinguished by its U308 concentration of high, medium, or low. Other ores contain substantial quantities of carbonaceous shale, typical of ore in the area. Feed is recovered from any one or combination of piles, in accordance with operating requirements and recovery factors. The run-of-mine ore, at 15% moisture, is transferred to a stationary, 24-inch by 24-inch grizzly. Undersize material falls into a 280-ton surge hopper located below grade, while oversize material is removed for preliminary size reduction. A sump-pump is located at the reclaim hopper to recover excess mositure and ore spillage for processing. Plant feed is continuously drawn from beneath the hopper at an average rate of 120 tons per hour by means of an apron feeder. The hopper is equipped with a low level safety system to protect the apron feeder assemble. The ore is transferred to a belt conveyor, elevated and discharged into a semi-autogeneous grinding (SAG) mill. Water addition at the discharge point is automatically controlled by a feed rate, moisture analyzer system located on the belt conveyor. GRINDING Minus 24-inch ore, at a rate of 120 mosit tons per hour and cyclone underflow at a rate of 137 moist tons per hour are combined with a controlled 186 gallons per minute fresh, warm plant process water. The mixture, at an average 68% solids, is fed to the SAG mill. The viscous mass is subjected to grinding at a temperature moderately above ambient resulting from use of the 140ºF well water. The 16.5'x5.0' Marcy Mill is driven by a 500 H.P. A.C. Motor. The mill will rotate at 73.4% of critical speed or 14.06 R.P.M. Eight percent of the mill's volume will be occupied by steel grinding balls; 22% by pulp. The mill is designed to rotate in either direction, in order to obtain maximum lifter and liner life. The SAG mill is equipped with a trommel having ½ -inch slots for removal of oversize material at its discharge and the large or oversize material is collected and either recirculated periodically or discarded. The undersize slurry is diluted to 64% solids with the addition of hot well water or mine water to the mill discharge sump. The flow is automatically controlled utilizing data from a gamma density meter. The slurry from the mill discharge sump is pumped to hydrocyclones to classify the particles. Underflow from the cyclones, at 80% + 28 mesh, is recycled to

Jan 1, 1979

By H. G. Doll

As a continuation of the earlier paper on the general subject of the SP log, a more complete analysis of certain features of the SP log in shaly sands is given. The pseudo-static SP in front of shaly sands is compared, on a theoretical basis, to the static SP in front of clean sands, as a function of the respective amount of shale and sand in the formation, and of the relative resistivities of the shale, of the uncontaminated part of the sand. and of the invaded zone of the sand. As a conclusion, the advantage of using reasonably conduc. tive mud in this case is shown. The discussion is illustrated by field examples. INTRODUCTION The discussion reported in the present paper is based on a theoretical analysis, and not on experiment. The field examples, joined to the text. are shown only as qualitative illustrations of the essential results of this analysis. Although the hypotheses made in the theoretical developments may perhaps be somewhat improved, it seems, nevertheless, that the results obtained account reasonably well for the actual phenomena, and give a fair approximation of their order of magnitude. The paper contains a mathematical analysis of a tri-dimen-sional distribution of potentials and current lines. due to spontaneous electromotive forces arising at the contact of shales and free electrolytes. as a function of the geometry and of the respective resistivities of the different media involved. It is assumed, although this hypothesis is not proven, that the emf's remain the same even if the shale occurs in very thin layers or in dispersed particles. It has already been pointed out 1,2,3 that, all other conditions being the same, the deflection of the SP log in front of a shaly sand is smaller than opposite a clean sand. When the thickness and the conductivity of a clean sand are large enough. the deflection of the SP log reaches a limiting value which is equal to the "static SP" of the clean sand. It is generally convenient to take the static SP of shale as the reference value or "base line." As a consequence, and for the sake of abbreviation, the expression. "static SP of a clean sand," is often used to designate the difference between the static SP of that sand and that of the shales, which difference is a measure of the total electromotive forces involved in the chain mud sand-shale. A similar limiting value is; also observed for the SP deflec-lion opposite a thick shaly sand, but it is smaller. just as if the total electromotive force involved were smaller in that case. This limiting value has been called the "Pseudo-Static SP" of the shaly sand. The static SP of a clean sand depends on the salinity of its connate water with respect to that of the mud, and, to a certain extent. on the differential pressure which controls the electro-filtration potentials, but it does not depend on the resistivity of the sand. On the contrary. the pseudo-static SP of a shaly sand depends not only on the salinity of its connate water and on the differential pressure, but also on the percentage of shale and on the resistivities of the shale, of the uncontaminated part of the sand, and of the zone invaded by the mud filtrate. If the three resistivities above were equal, the pseudo-static SP would be proportional to the percentage of sand in the shaly sand, and its departure from the static SP of a clean sand having the same connate water would simply be proportional to the percentage of shale. In that case, the pseudo-static SP of a shaly sand containing 10 per cent of shale would he 10 per cent less than the static SP of a clean sand. When. however. the sand is. on the average. substantially more resistive than the shale. the percentage of departure of the pseudo-static SP from the static SP of a clean sand is much larger than the percentage of shale. For that reason, the peaks of the SP log opposite shaly sands are systematically of smaller amplitude when the sands are oil-bearing than when they are water-bearing, all other conditions being the same. This feature is observed even when the sand beds are thick. and even when they do not contain a large percentage of shale. All this has already been described in all earlier publication", but mostly in a qualitative way. The present paper will analyze in more detail the action of the local SP currents which are generated inside of the shaly sands, and which are responsible for the abnormally low value of the pseudo-static SP. The quantitative computations have been extended to the general case of thin interbedded layers of sand and shale, where the resistivities of the shale and sand streaks do not have the same value: they are summarized in charts giving values of the pseudo-static SP of a shaly sand as a function of the different parameters involved. DEFINITIONS The static SP of a clean sand has been defined as the potential that would exist in the mud opposite that sand, were the SP current prevented from flowing. Such an ideal condition is represented on Fig. I-A. By analogy, the pseudo-static SP of a shaly sand can be defined as the potential that would exist in the hole, if the circuit shaly sand — surrounding shales — mud column were interrupted by the insulating plugs placed at the boundaries

Jan 1, 1950

By C. S. Giggins, F. S. Pettit

The oxidation properties of Ni-Cr alloys with fine grains, coarse grains, and deformed surface layers have been studied at temperatures of 900" and 1100°C in 0.1 atm of oxygen. The oxidation rates of alloys containing between 10 and 30 wt pct Cr have been found to be dependent upon the grain size of the alloy. Finegrained alloys had smaller oxidation rates than coarsegvained alloys because of the selective oxidation of chromium at alloy grain boundaries. In this compositional range alloys with deformed surface layers behaved similar to fine-grained alloys due to recrys-tallization of the deformed surface layer. In the preceding paper1 it was found that during the oxidation of Ni-Cr alloys, the volume fraction of precipitated Cr2O3 could be greater at alloy grain boundaries than at other areas of the alloy surface. In the case of alloys with chromium concentrations equal to or greater than 30 pct,* the volume fraction of Cr2O3 *AIL compositions are given as weight percent unless specified otherwise. precipitated at grain boundaries and within grains on the alloy surface both exceeded the critical amount required for lateral growth of the Cr2O3 particles and the surfaces of these alloys were completely covered with a continuous, external layer of Cr203 during oxidation. However, in the case of alloys with chromium concentrations between approximately 5 to 30 pct, the volume fraction of precipitated Cr2O3 exceeded the critical value required for external scale formation only at grain boundaries.but not within the interior of the grains. Consequently, the surfaces of these alloys had external scales of Cr3O3 over the grain boundaries but internal Cr2O3 subscales with external scales of NiO away from the grain boundaries. Under these latter conditions, it was found that chromium could diffuse laterally in the alloy from those areas covered with an external layer of Cr2O3, i.e., grain boundaries, to areas where the Cr2O3 was present as a subscale. This diffusion of chromium resulted in an increase in the volume fraction of Cr2O3 precipitated in the sub-scale zone and continuous layers of Cr2O3 could be formed at the subscale front in these regions. For the alloys used in the previous studies,' continuous layers of Cr2O3 were formed on Ni-20Cr alloys in the subscale regions after approximately 30 hr of oxidation at 900°C. For shorter periods of oxidation, the Cr2O3 layer was semicontinuous with the continuous portion at the subscale front emanating from points where the Cr2O3 had been formed as an external scale over alloy grain boundaries. Some lateral growth of a Cr2O3 layer in the subscale region was observed on Ni-15Cr and even Ni-10Cr alloys but this layer was never continuous after 30 hr of oxidation. These results indicate that the selective oxidation of chromium in Ni-Cr alloys with chromium contents between 5 to 30 pct may be dependent upon the grain size of the alloy. Fine-grained specimens in this compositional range should have a larger fraction of the surface covered with external Cr2O3 than coarsegrained specimens and the subscale areas required to be sealed via lateral diffusion of chromium should be smaller. It is therefore to be expected that a continuous layer of Cr2O3 can be formed on alloys in this compositional range after short periods of oxidation providing the alloy grain-size is sufficiently small. bo studieS2,3 have established that the oxidation behavior of alloys can be significantly influenced by pretreatments which produce mechanically deformed surfaces. It has been found that deformed surfaces usually promote the selective oxidation of elements in alloys and it is believed that these effects are due to rapid diffusion of elements in the deformed layer. In view of the previous results,' which showed that alloy grain boundaries may play an important role in the selective oxidation of chromium in Ni-Cr alloys, deformed surfaces may promote the selective oxidation of elements in alloys as a result of the numerous grain boundaries formed on the alloy surfaces via recrys-tallization during heating to the oxidation temperature. The purpose of the present studies was to determine the effect of alloy grain size and surface deformation on the selective oxidation of chromium in Ni-Cr alloys at temperatures of 900" and 1100°C in 0.1 atm of oxygen. EXPERIMENTAL The average grain diameter of the alloys used in the previous studies1 was not less than 0.04 mm and alloys with compositions between 5 and 30 pct chromium had average grain diameters between 0.04 and 0.14 mm. Since the oxidation kinetics were already available for these relatively coarse-grained alloys, it was desirable to use these same alloys in the present studies. The surfaces of the alloys listed in Table I of the previous paper were deformed by using a Model F S.S. White Industrial Airbrasive Unit, which delivered a controlled mixture of 25 µ Al2O3 particles in a stream of dry air at high velocity against the surface of a specimen. The amount of surface deformation produced by this treatment was not determined but a re-crystallized layer about 15 µ thick was formed upon annealing deformed specimens. The grain size at the surface of the specimens was reduced to an average grain diameter of 0.01 mm by annealing the deformed specimens, i.e., grit-blasted,

Jan 1, 1970

By M. J. Marcinkowski, Gordon E. Lakso

An extensive investigation has been made of the deformation behavior associated with the Fe3Si super-lattice using transmission electron microscopy techniques. Above 243°K the stress-strain curve exhibits three stages. Stage I occurs at a very low stress level and is related to the generation of perfect superlat-tice dislocations. Stage II is characterized by an extremely rapid rate of work hardening and is associated with the Taylor type locking of these superlattice dislocations. Finally Stage III is related to dynamic recovery processes since the work hardening rate is very small. Below 243ºK, only Stage I is observed, but it occurs at a much higher stress level. This latter observation is related to the generation of imperfect dislocations in Stage I with the consequent production of second nearest neighbor antiphase boundaries. The reason for this is that insufficient thermal energy is available at these low temperatures to generate the complete and perfect superlattice dislocations. It has been shown that the fully ordered FeCo alloys, i.e., those possessing the B2 type structure, exhibit three distinct stages of work hardening whereas the corresponding disordered alloys show only one.&apos;" This difference in behavior between the disordered and ordered alloys has been attributed to the fact that dislocations in the former case travel only as ordinary 1/2ao(111) types whereas in the latter case the move through the lattice as coupled 1/2a0(111) dislocations separated by an antiphase boundary (APB), i.e., the so-called superlattice dislocation. Although some preliminary work has been carried out concerning plastic deformation in ordered alloys possessing the DO3 type superlattice,3 no detailed analysis similar to that described in Refs. 1 and 2 has been attempted. Specifically, it has been suggested that the superlattice dislocation in this particular type structure should consist of four ordinary 1/2ao<111> types bound together by first and second nearest-neighbor APB&apos;s. Fe3A1 and Fe3Si are the two classic alloys possessing the DO3 type lattice; however, because of the somewhat higher ordering energies associated with the FesSi alloy, which in turn assures that dislocations will travel through the lattice as perfect superlattice dislocations under at least some conditions, it was chosen for the present investigation. Because of the extreme brittleness of Fe3Si, all deformation was done in compression. Stress-strain curves were obtained using both polycrystalline samples as well as single crystals. In the latter case the crystals were oriented so that deformation could be controlled either by single or double slip. They were then wafered parallel to and at various angles to the operative slip planes. These wafers were in turn examined by transmission electron microscopy (TEM) techniques in order to determine the extent of the interaction from the dislocation configuration contained therein. EXPERIMENTAL PROCEDURE The alloys used in this investigation were arc melted under helium from electrolytic iron of greater than 99.90 wt pct purity and transistor grade silicon of 99.99 wt pct purity. A typical analysis of interstitial impurities showed 120 ppm 0, 15 ppm N, and 65 ppm C Because of the extremely low ductility of the Fe3Si alloys, it was necessary to spark cut 0.230-in. diam polycrystalline cylinders 0.400 in. long from arc-melted fingers using a thin-walled brass tube as a cutting tool. The polycrystalline alloys could not be recrystallized since very little strain was induced in preparation. However they were annealed at 1273°C for 15 min in evacuated vycor capsules to relieve any cooling stresses that may have developed during solidification and then air cooled. The resulting grain size of the alloy was 0.50 mm. According to warlimont4 1273ºC is just within the single phase field where FesSi possesses the DO3 type lattice. In addition because of this high critical ordering tem-ature, air cooling from this temperature was believed sufficient to fully order all of the Fe3Si samples used in the present investigation. For the same reason, no attempt was made to achieve any degree of disorder by quenching. In fact, rapid quenching from 1123°K caused cracking. Such cracking was first suggested by sato5 with respect to the experimental observations of Glaser and Ivanick.6 Single crystal compression specimens were spark cut from single crystal ingots grown in a Bridgman type furnace. The iron and silicon for the crystals was prealloyed by arc-melting two 130-g buttons which were cut into small pieces before remelting in the furnace. This procedure resulted in a long-range inhomogeneity of 0.5 at. pct Si between the top and bottom of the 2-in.-long single crystal ingot, which was assumed to be negligible in the present investigation. The single crystals, after orienting and spark-cutting, were about 0.37 in. by 0.37 in. in cross section and about 0.5 in. long. True stress-strain curves were obtained using an Instron Tensile Testing machine in conjunction with techniques described previously. 1,7 The strain rate was 0.05 in. per in. per min. Prior to testing, the ends of all the compression cylinders were hand polished using a special jig to insure parallelism after which the sides of the samples were electrochemically polished to eliminate stress risers and to facilitate slip line observations. Test temperatures between 77" and 823°K were obtained using various cooling and heating media as described in Ref. 7 while at the upper end of this temperature range, a mixture of equal

Jan 1, 1970

By W. A. Turcotte, A. D. Paananen

Introduction The large reserve of fine grained oxidized iron-formation at the Tilden mine has been the object of research and development efforts to concentrate the iron oxides as far back as 1949. Due to the nonmagnetic nature of the ore and the fine grinding required to liberate the iron oxide minerals, this crude ore was not amenable to concentration by conventional methods. The iron oxides of the Tilden, ore body have a grain size of less than 25 microns and recovery of the finer, well-liberated iron oxides is essential. Conventional methods of desliming employing cyclones or thickeners were not feasible because of the excessive loss of iron oxides in the finer fractions. Development of selective flocculation-desliming was a key to commercialization of the process. Operations started in late 1974 with Algoma Steel Corp. Ltd., J & L Steel Corp., The Steel Company of Canada Ltd., Wheeling-Pittsburgh Steel Corp., Sharon Steel Corp., and The Cleveland-Cliffs Iron Co. as participants. Cleveland-Cliffs operates and manages the operation. Development of the Tilden Flowsheet The geology and ore reserves of the Tilden mine have been detailed in a paper by Villar and Dawe (1975). A joint program was undertaken in 1961 with the US Bureau of Mines in Minneapolis using the flowsheet developed by the Bureau employing the selective flocculation-desliming and calcium activated anionic silica flotation method (Frommer, et al, 1966; Frommer, 1964; Frommer, Wasson, and Veith, 1973). During this time, parallel testing at Cleveland-Cliffs Research Laboratory and Pilot Plant centered on the same type of desliming but was followed by the cationic flotation of silica with amine collectors (Columbo and Jacobs. 1976). The cationic silica flotation system was eventually chosen for its overall efficiency and simplicity. Regardless of the flotation method chosen, the technique of selective flocculation-desliming prior to flotation is the key to the success of the process. The flowsheet is described in detail by Villar and Dawe (1975). [Figure 1] shows a simplified one-line flowsheet of the Tilden concentrator. A total tailings thickener has been added to the original flowsheet and was placed in operation in 1978. The total-tailings thickener overflow reports to the reuse water pond and the underflow is pumped approximately 8 km (5 miles) to a storage basin. A flowsheet of the reuse water system is shown in [Fig. 2]. Selective Flocculation-Desliming Data have been published on the mechanisms and factors affecting selective flocculation in iron oxide-silica systems. The intent of this paper is not to discuss the theoretical aspects of selective flocculation, but rather to present experience gained from the commercial Tilden operation and from bench and pilot plant testing of fine-grained oxidized iron ores. From the bench and pilot plant testing prior to plant startup, certain reagent combinations and rates for the commercial Tilden plant were established. In the experience gained from three years of plant operation at Tilden, some of these reagent dosage rates have required significant adjustments due to changes in reuse water quality and to meet the requirements of varying ore types. Reuse Water The process water quality is a major concern at the Tilden mine and is constantly being monitored for selected chemical and physical characteristics. This monitoring has continued on a regular basis in order to gain a more thorough understanding of the interactions taking place in a dynamic water system and particularly as water quality is influenced by seasonal variations. Control of the reuse water chemistry is essential to the Tilden process both in the selective flocculation-desliming and flotation stages of concentration. With roughly 75% of the reuse water used in grinding-desliming operations, it is readily apparent that the biggest "reagent" in the selective flocculation-desliming process is water. Not enough can be said about the close control that must be exercised on the overall reuse water system. Control of the chemical treatment of the feed to the total tailings thickener is of utmost importance in order to produce a reuse water for the concentrator that is compatible with all stages of the concentrating process. There are many analyses made which aid in judging the quality of the water. Some of these are shown in [Table 1]. Five are particularly important and are monitored daily so that reagent adjustments can be made as required: suspended solids, calcium hardness, pH, dissolved silica concentration and temperature.

Jan 1, 1981