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By D. J. Carney, R. L. Stephenson

M. O. Holowaty (Inland Steel Co., Chicago)—Messrs. Stephenson and Carney should be commended on the presentation of this very interesting paper. They have introduced new ideas on sampling of sinter and have thereby made a valuable contribution to the qualitative evaluation of sinter. The paper undoubtedly presents a novel approach to the problem of sinter sampling and this is to be regarded as its main value. Sampling of sinter from the pallets prior to the discharge is by no means new. This sampling procedure was used in the past by Philbrook; Hamilton and Ameen and others. The sampling basket used by the authors of this paper, however, permits one actually to obtain a good sample of the sintering mixture. The basket which can also be described as a removable pallet goes then through the entire process of sintering and thus offers a sample which is positively representative of the batch in question. The authors' method of evaluating physical properties of sinter seems to be not altogether clear, however. It is felt that the following points should be clarified: I—In the section describing the shatter test the authors state that a 5-lb sample was used in the actual tests. There is no mention of the initial size of sinter particles used and the method of obtaining this 5-lb sample from the sinter cake weighing about 100 to 120 lb. It is also surprising that the sample was dropped from a height of 3 ft instead of the conventionally used 6 ft. Unfortunately there is no mention of the number of drops to which the sample was subjected. 2—The tumbler test used by Messrs. Stephenson and Carney in testing the samples of iron ore sinter was performed in a standard ASTM coke tumbling drum. The drum measures 3 ft in diam and is equipped with two 3-in. lifter bars. The distance of fall in this testing drum is identical with the distance used in the previously described shatter test. The number of drops in the tumbling drum is approximately double the number of revolutions. The basic difference then between the drop-shatter test used by Stephenson and Carney and the tumbling test seems to be the number of drops to which the sample is subjected and the abrasive action to which the sinter particles are exposed in the rotating drum. In any case, the choice between the shatter test and the tumbler test seems to be only a matter of taste and convenience. 3—The authors of the paper state that "in order to insure getting results that were representative . . . the entire sample obtained by means of the sampling basket was used as the starting material for the tumbler test. If the sample broke in pieces, all the pieces were used regardless of their size." The initial size of the material is known to be a very important factor, influencing seriously the final size of the particles. This should be particularly important in case of hard sinters. It seems also very troublesome and inconvenient to charge 12-in. cubes of sinter weighing about 100 lb into the testing drum. This could be eliminated by taking a standard fraction of somewhat smaller particles.

Jan 1, 1954

By J. F. Shirley, H. K. Burke

In January, 1964, three years of laboratory and pilot plant test work reached its culmination. The San Manuel Div. of the Magma Copper Co. switched its molybdenite extraction circuit from the sodium hypochlorite-ferrocyanide process to the new process described in this paper. The San Manuel Div. Concentrator of the Magma Copper Co. is located near San Manuel, Ariz., approximately 45 miles northeast of Tucson. The concentrator treats approximately 35,000 tons of ore per day containing about 0.85% copper and 0.025% molybdenite. The San Manuel ore body is a disseminated copper deposit in quartz monzonite and quartz monzonite porphyry. The primary copper mineral is chalcopyrite; minor amounts of chalcocite, bornite and covellite are also present in varying degrees. Pyrite makes up somewhat more than half the sulfide content of the rock. The molybdenite is nearly always found on fracture faces or associated with quartz veinlets.' On January 9, 1964, the sodium hypochlorite-ferrocyanide process that had been used for molybdenite recovery was replaced by the following process: 1) A conditioning step with hydrogen peroxide, sulfuric acid, sodium cyanide and zinc sulfate. 2) A two stage rougher flotation step with stove oil and sodium ferrocyanide. 3) One cleaning step with sodium ferrocyanide. 4) One cleaning step with sodium hypochlorite and potassium ferricyanide. 5) Four cleaning steps with an anti-foam Exfoam 636 and potassium ferricyanide. COPPER CONCENTRATOR Since molybdenite recovery is closely allied with previous steps in the operation, the copper concentrator will be briefly described first. The ore is ground to approximately 5% plus 65-mesh and 65% minus 200-mesh in a two-step grinding circuit consisting of rod mills in open circuit and ball mills in closed circuit with rake classifiers. A straight forward rougher circuit producing only one concentrate is used. The rougher concentrate is combined with the discharge from the regrind ball mills and is then cycloned, with the overflow going directly to the cleaner circuit and the underflow going to the regrind mills. The feed to the cleaner circuit is approximately 90% to 94% minus 325 mesh. The cleaner circuit consists of two stages in which the tailing from the final cleaner stage is returned to the primary cleaner and the primary cleaner tailing is returned to the rod mill discharge at the head of the circuit. Approximately 0.013 lb of the allyl ester of amyl xanthate (S-3302) per ton of ore is added to the rod mills and approximately 0.006 lb of sodium isopropyl xanthate (2-11) per ton of ore is stage added down the rougher banks. The frother used is methyl iso-butyl carbinol (MIBC) and it is added at various points in the grinding and the rougher flotation circuits. Lime is added to the ball mills and/or to the rod mills to maintain the desired pH in the rougher flotation. The rougher flotation has been operated at various pH levels between 10.5 and 11.5. Fuel oil is added in the regrind mills at a loading of approximately 0.01 lb per ton of ore. Factors Which Directly Affect Molybdenite Recovery: It has been found at San Manuel that, to obtain maximum molybdenite recovery, an oily type flotation collector must be added in the grinding circuit. When molybdenite is crystallized in the characteristic platy form, which appears to be the case almost all of the time at San Manuel, its surface has a greater tendency to adhere to oils than water. 2 Of the several oily type collectors available on the market, the allyl ester of amyl xanthate seems to be the best molybdenite promoter. The floatability of San Manuel's molybdenite does not appear to be affected, at least up to a pH of 11.5, by lime. It has been reported that at several plant operations high lime alkaline circuits have adversely affected the molybdenite recovery.3,4 It has been found through microscopic studies that the bulk of the chalcopyrite will float before the molybdenite in both the rougher and cleaner circuits. It is necessary to add fuel oil to the regrind mills to enhance the

Jan 1, 1965

By W. E. Horst

Flotation feed was hell constant for batch laboratory tests and continuous pilot and commercial plant tests on a spodumene ore to determine the relation.ship between flotation behavior and the size and type of flotation machine. Results. showed that additional retention time would be needed in the plant flotation circuit to duplicate laboratory or pilot plant performance when operating Dariables of the. plant flotation machines remained unchanged. During the past few years of operation at Kings Mountain, N. C., full-scale flotation has generally yielded poorer metallurgical results than those obtained in the laboratory or pilot plant. After 2 min in each size of Denver flotation machine in the spodumene rougher circuit, pilot recovery (41.7 pct) was almost three times that of the plant (13.6 pct) and about twice that of the laboratory (21.0 pct). Pilot and commercial plant data are based on continuous operation, and laboratory results on batch testing of rougher performance. The ore is a pegmatite containing spodumene, feldspar, quartz, mica, and amphibolite. Data were accumulated to evaluate the scale-up relationships between laboratory, pilot plant, and plant flotation cells. Flotation rate constants and dimensional analysis were used to define the relationship existing between flotation behavior and scale of operation. Measuring Flotation Rates: Efficiency of the rougher stage was measured by flotation rate, defined as the recovery that can be attained within a given time." • The problem of measuring flotation rates has been studied by many investigators, whose methods have varied. Published studies an flotation kinetics have been summarized recently by Gaudin.l When comparing flotation rates, it is more satisfactory to compare straight line relationships between recovery and flotation time than exponential curves resulting from plotting cumulative recovery vs cumulative flotation time. To convert the exponential curves into straight line relationships, rate equations were applied similar to those used in chemical kinetics to study the velocity of chemical reactions. Two rate equations have been used in this investigation to measure efficiency of the various cells: 1.5 order Rate Equation dR/dt=k (A-R)1.0 [Eq.1] Second Order Rate Equation dR ---- - K(A - R)2 [Eq. 2] dt Upon integration Eq. 1 becomes 1/(A-R)1/2=k/2 t+1/A1/2 [Eq. 3] and Eq 2 becomes: or fEq. 4] t/R=1/A2K+t/A [Eq. 4] where R = cumulative recovery [percent recovery of total value (spodumene) in flotation feed]. A = initial concentration, pct (100 pct of mineral to be separated). t = flotation time (minutes). k = flotation rate constant for 1.5 order rate equation [units: 1/(percent) 0.5 (min)1. K = flotation rate constant for second order rate equation [units: 1/(percent) (min) 1.

Jan 1, 1959

By J. Walter Snavely

PRESENT-DAY processing in many industries, calcining, sintering, briquetting, beneficiation and nodulizing, increasingly calls for the handling of large volumes of hot bulk materials. Various types of conveyors have been employed. This discussion will cover the factors governing their selection. For temperature ranges up to 400°F, or approximately 200 °C, a wide range of conveyors is available. Special constructions of rubber conveyor belts, steel conveyor belts, vibrating and shaker conveyors, apron conveyors, and drag chain conveyors, all are used successfully. As temperatures go well above 400 2F, however, choice of conveyors is narrowly limited. This paper will consider the problem of handling bulk materials only where the temperatures exceed 400°F. The arbitrary selection of 400 °F as a dividing point undoubtedly can be challenged, as special conveyor belting constructions are available which are suitable for temperatures in excess of 400°F. However, when the relatively short life of such belts and the cost of their replacement, with the attendant down time, are balanced against the reliability and long service life of the properly designed steel constructed units to be discussed, there is little question in any operator's mind that the special belts are more expensive to use. Because the conveyors under study are for the handling of bulk materials, inevitably including a high proportion of fines, obviously wire mesh belts cannot be included for consideration. Even though this type of conveyor is widely used at high temperatures, i.e., for carrying glassware through a lehr, it is unsuited for the conveying of bulk materials, and therefore will be excluded from further discussion in this paper. Preliminary to the study of the conveyor itself is the determination as to whether the material is to be cooled while it is being handled, or whether the processing requires retention of all heat and the maintenance of a given temperature range. In the majority of cases cooling is incidental to or part of the handling process, when the handling, for example, follows completion of sintering, roasting, calcining, refining, or some other process. To meet such operating conditions successfully, the conveying medium used must have: 1—a construction capable of withstanding maximum initial temperatures of the material being handled. 2—a construction providing efficient heat transfer for cooling. 3—a construction providing dependable operation and long life with minimum service requirements, and 4—a construction providing controlled and efficient conveying. Under the usual conditions of cooling during the handling, the construction selected to withstand the initial maximum temperatures does not necessarily involve using alloys, as excellent results can be achieved with normal carbon steels and cast irons, when they are properly applied and proportioned. The earliest and simplest type of conveyor for handling very hot materials is the cast steel drag chain conveyor, still widely used for handling hot cement clinker, as illustrated by Figs. I and 2. Because of the rugged and generous proportions of the chain link design, low carbon steels are entirely suitable for the links. The pins, however, must be alloy steel. The simple, rugged construction of this type of conveyor makes it readily capable of withstanding high initial temperatures, even though the chain is operating buried in the material. The drag-chain type of conveyor has advantages and limitations. Although the efficiency of the heat transfer is relatively poor, the life of the conveyor is reasonably long, and because of its crude simplicity it does not require much servicing. However, as a conveyor, it is limited in capacity, and largely limited to horizontal runs. Furthermore, because of the crude design, heavy weight, and the chain operating at the temperature of the material, greatly reducing permissible operating chain pulls, this type of conveyor is limited to relatively short centers. Another type of conveyor that has been used for very hot materials is the cast pan conveyor. Because of its very generous proportions the cast pan, which is made of either cast iron or malleable iron, can withstand initial maximum temperatures. It also provides efficient heat transfer for cooling. Further, it is on efficient conveyor construction, which can be used for inclines. Because the chain employs rolling friction instead of sliding friction, and is not in the maximum temperature zone, much longer centers are possible. It is this type of conveyor that is frequently used in the casting of various metal pigs, pig iron, and aluminum; it is obvious, therefore, that very high initial temperatures are being handled. With this kind of conveyor the return run is frequently sprayed with water to accelerate heat transfer. The build-up of residual heat in the very heavy cast pans is thus overcome. The outboard roller steel pan conveyor is an improved pan conveyor' which provides high rates of heat transfer and substitutes formed steel pans for the heavy cast pans. It is a very efficient conveying medium. The details of this particular construction are clearly shown in Fig. 3. An early application of this type of conveyor is shown in Fig. 4. In this case the conveyor units are handling roasted phosphate rock at average temperatures of 1000" to 1500°F, and frequent maximum temperatures as high as 1900°F. Several widths are used. The capacity of the unit at a speed of 50 fpm is approximately 30 tph per inch of width at peak loadings, average capacity being about 1/3 of peak loading. The assembled conveyor is shown in Fig. 5, with views of both the top and the underside to show all the construction details. In particular, the following general design principles were carried out in this construction:

Jan 1, 1954

By G. Gutierrez, M. C. Fuerstenau, J. D. Miller

The response of pure goethite and two natural iron ores to flotation with potassium octyl hydroxamate is presented. The ores contained 12.5% and 39.7% iron; concentrates containing 69.5% and 61.5% iron, respectively, were obtained with additions of 0.4 lb/ton hydroxamate. Selective flotation of iron oxides from associated gangue minerals has proven difficult in the past with standard collectors, such as sulfonates and fatty acids. As a result, recent work has been directed toward silica flotation with calcium activation after the iron oxide surfaces have been blinded with certain starches.& The results obtained with this technique are encouraging. This method, however, involves depressing a mineral that would normally float under the conditions employed. A preferable method would involve floating the iron oxide because quartz is naturally depressed until activated. Earlier work involved this approach, but in view of the results obtained, collectors that exhibit greater specificity for iron are needed. During an investigation of chrysocolla flotation with potassium octyl hydroxamate, a metal chelate former, it was observed that pyrite floated well with low additions of collector.3 This phenomenon suggests that hydroxamate has a rather strong affinity for iron, and that other iron-bearing minerals should also respond to flotation with octyl hydroxamate. The object of this paper is to determine this response. The experimental work was divided into areas, the first of which involved microflotation experiments with pure goethite with an apparatus described previously.4 Goethite was selected because of the ease

Jan 1, 1968

By A. Chakravarti, C. C. Harris

First-order kinetics adequately describe the recovery of a single floatable species. The conditions under which cell behazior can be simulated by a single-phase model are discussed; deviations from the multiphase model are accountable. The rate coefficient depends upon operating conditions besides particle size and mineralogy and the physicochemical properties of the pulp constituents. The model for a real flotation system combines the first -order, single-phase model with a distribution of noninteracting species. Methods for determining the distributions are discussed. Applications of the procedure to flotation studies are outlined. The mathematical description of flotation as a rate process can be traced to Garcia1 who reported cases in which the fraction of solid remaining in a cell in semi-batch operation, F (for Notation used throughout, see Table l), decays exponentially with experimental time T. F = e-k>r. (1 The differential equation leading to Eq. 1 is dF ----= -k,F (2) dT which is recognized as the algebraic analog of the chemical kinetics first-order reaction equation.' This has led to suggested mechanisms for flotation and mathematical relationships other than first order: with very little foundation other than analogy. The flotation rate coefficient kl can be accounted for as the product of several probabilities concerning bub-ble-particle attachment and detachment, aggregate entry from pulp into froth and return from froth to pulp, and removal from froth as concentrate. These factors clearly contribute to the magnitude of the overall flotation rate, and there are others also, but describing flotation merely as the overall probability of consecutive gross events hardly appears a significant advance in understanding. More recently, it has been described in terms of a stochastic process, which promises to provide a more satisfactory approach? The present paper will be more concerned with a phenomeno-logical description of the flotation process than with an account of the internal mechanics of the process. This is not a statement of priorities, or a judgement of the relative importance of these options but rather an assessment of what is presently possible. Previous investigators'" have developed methods for interpreting semi-batch flotation data in terms of species distribution. These methods have been used in an attempt to explain flotation behavior in response to changes in operating conditions." Inspection of examples given in the literature" reveals discrepancies between theory and experiment, and data obtained in the Colum-

Jan 1, 1971

By Norman F. Schulz

The technical excellence of separation achieved in a mineral concentration process, or any other process where two constituents of any kind are physically separated from each other, is expressed uniquely and quantitatively by the separation efficiency (E,): E. = (R - R,), where R is the percentage of the valuable constituent and R, the percentage of the waste appearing in the concentrate. The separation efficiency meets all the usual requirements of a true efficiency and is independent of the nature of the process used. It can be visualized as the percentage of the feed material that, in effect, actually undergoes complete separation while the rest of the feed is distributed randomly and unchanged into the two separation products. The problem of quantitatively assessing the results of a mineral concentration process recurs whenever one tries to decide which of two similar test results is more desirable. Although economic factors may greatly influence the eventual choice of process, they usually have very little to do with the technical excellence of a separation and are, therefore, not a consideration at this point. Concentrate grade and recovery, used simultaneously are the most widely accepted measures of performance, and if both the grade and the recovery are greater for one case than for another, the choice is simple. However, when the results of closely competitive operations are compared, it often occurs that one set of test results shows a higher grade and a lower recovery than another and the choice is no longer obvious. Laboratory investigations of a mineral concentration process applied to a given feed material can often be designed to yield successive paired values of concentrate grade and recovery over the practical range of the course of a separation process.' These results may be plotted as some combination of the cumulative values of grade, weight recovery, or unit recovery in the concentrate. The resulting correlations for two slightly different sets of conditions can then be compared qualitatively, or even quantitatively in particular cases.2 Such comparisons are not applicable, however, to individual concentration test results. The reduction of a separation test result to a single efficiency figure, preferably as a percentage of ideality, would afford the simplest and most precise means for comparing the results of competitive tests. The effective proportion of feed material that undergoes complete separation in any real concentration process, hereafter referred to as "separation efficiency," is such a figure. It is essily calculated from the usual feed and product data and can even be evaluated, in many cases, from data in publications describing various concentration processes. Mineral Concentration Processes and Their Evaluation Mineral Concentration Processes: The overall concentration process may be subdivided into a liberation process which severs the bonds between unlike mineral species, and a separation process, which sorts the resulting liberated particles into two products—a concentrate rich in mineral A and a tailing lean in mineral A. A perfect mineral concentration process is represented by the idealized scheme shown in Fig. 1 in which liberation and separation are not differentiated. The feed mixture consists of two mineral types which are completely liberated early in the process and then sorted by type. All the mineral particles of type A which respond uniquely to the selection forces acting in the ideal separator, report to the concentrate, while all the mineral particles of type non-A report to the tail-

Jan 1, 1971

By A. H. Tousley, C. L. Vickers

The author presents a history of the development of the Port of Long Beach from its beginning in 1905 to the present construction of a new bulk loading facility. In describing this new facility, emphasis is placed on such aspects as rail car and truck handling, gravity yard, rotary car dumping, bottom dumping, car dumping equipment, feeding and conveying, ship-loader-stockpiler, stockpiling and reclaim, dust control, and construction problems. Development of the Port of Long Beach, Calif. (see Fig. I), began in 1905. In that year private Long Beach interests acquired 800 acres of marshland in the Inner Harbor area and initiated a program of channel dredging and land reclamation. Efforts to establish a municipal port began a short time later and culminated in the opening of Pier 1 in 1911. An event of even greater significance to the port future also occurred in 1911. This was the tideland grant enactment by the state legislature that conveyed to the city of Long Beach all of the state's rights in the tidelands located within the city boundaries. The grant stipulated that all revenue produced from the tidelands be used for the development of navigation, commerce, and fisheries. During the next 15 years up to 1926, the port's commercial growth was steady, although relatively slow. During this period, several fundamental steps in the development of a major harbor were accomplished. These included the creation of the Board of Harbor Commissioners in 1917, the dredging of the Inner Harbor and entrance channel to accommodate deep draft vessels, and the correction of a serious silting problem through construction of the Los Angeles River Flood Control Channel. These efforts bore fruit during the ensuing 10 years in the establishment of more than $30,000,000 worth of industrial plants and private marine terminals in the Inner Harbor. Commerce over municipal wharves also increased more rapidly, and it became necessary to start the development of the harbor in a seaward direction. In 1936, oil was discovered in the harbor area, and the first Harbor Dept. well was brought in 2 years later. It soon became evident that the city of Long Beach would receive substantial revenue that, under the terms of the state grant of 1911, must be used in the development of the port. It was further realized that to properly guide and coordinate such development, a long range comprehensive plan was essential. The Board of Harbor Commissions contracted with G. F. Nicholson and J. F. Collins, consulting engineers, for the preparation of such a report that was completed in 1940. The Nicholson-Collins report proposed the development of an East Basin, between the entrance channel and the Los Angeles River, and a West Basin in the area now occupied by the Long Beach Naval Shipyard. Twenty full-size berths were to be provided in the East Basin and more than fifty berths in the West Basin. The report recommended extensive modern terminal facilities for handling waterborne cargo. Attention was given to improvement of the street and highway system to give better vehicular access and to avoid congestion within the port. To a large extent the recommendations of the Nicholson-Collins report were followed in the subsequent development of the East Basin. This development was greatly accelerated by World War 11, with its need for increased oil production and additional shipping facilities. The wartime and immediate postwar construction was largely concentrated on Pier A, with the development of Piers B, C and D following in more recent years. An important factor in the planning of these piers was their dual function as oil drilling sites and cargo piers. The Navy's acquisition of the West Basin area made it necessary to depart from the Nicholson-Collins plan and find other areas for future expansion. At the present time only one Inner Harbor Port-owned area having protected deep water frontage remains undeveloped. This is the property fronting on Channel 2 in the Inner Harbor acquired in December 1957, from Pacific Dock and Terminal Co. The growth of the port has been in a southerly direction since the construction of the Long Beach Naval Shipyard in the West Basin area. Recent extension of Piers F, G and J into the ocean follows typical port construction methods developed during the expansion of the Middle Harbor during the last 20 years.

Jan 1, 1963

By H. W. Hitzrot, R. H. Limons, L. V. Fegan

The testing of a new method for measuring pellet-bed temperatures in the traveling-grate or shaft furnace is presented in this paper. This method, using silver-palladium alloy wires as indicators, overcomes disadvantages existing in the methods based on the use of thermocouples which are costly and time-consuming. The authors have developed and tested a new method for measuring pellet-bed temperatures in the traveling-grate or shaft turnace through the use of silver-palladium alloy wires as indicators. This method overcomes the disadvantages of existing methods, which, based on the use of thermocouples, are generally costly and time-consuming. For example, in the case of the traveling-grate furnace, the installation of an instrumented pellet, together with its attached thermocouples, requires stopping the furnace, with resulting loss of production time. As for temperature measurement in the shaft furnace by the existing method, the movement of the shuttle-car feeder seriously interferes with both the actual insertion and the resulting distribution of continuously fed thermocouples. The authors have overcome these difficulties by using, as temperature indicators, a range of silver-palladium alloy wires that can be easily recovered, after firing at any desired location in a pellet furnace. Silver-palladium wires having specific melting points from 2000°F to 2500°F are enclosed in ceramic capsules. Each capsule is placed in a ni-chrome wire basket filled with a sample of green pellets, the baskets being the means for recovering the capsule after firing. After the encapsulated wires have been exposed to the firing cycle, the alloy wires are examined to determine the maximum temperatures attained by the pellet sample. This method gives accurate, rapid determinations of maximum temperatures attained during pellet induration in either the traveling-grate or shaft furnace The new method has the following advantages over existing methods: 1) it is less expensive; 2) no furnace downtime is required; 3) the testing units can, be made in quantity in advance; 4) measurements can be made in rapid succession at all desired locations; 5) measurements of maximum firing temperatures are more accurate; 6) on the basis of points 4 and 5, the new method gives a more reliable picture of the correlation of pellet quality, petrographic textures and maximum temperatures, thus providing a more effective quality-control tool. A disadvantage of the method is, of course, that the measurements are not continuous and only maximum temperatures are determined. TEST PROGRAM Preparation of Materials: CALIBRATION OF ALLOYS - In the following discussion the term "bead point" (see Fig. 1) refers to the temperature at which each alloy begins to melt and form a bead. Fig. 2 gives the bead point curve for silver-palladium alloys. Before the alloys are used in plant tests, the bead point of each alloy should be determined in the laboratory to verify the alloy compositions. Once a given alloy has been calibrated, it can be

Jan 1, 1965

By D. W. Fuerstenau, G. D. Gumtz

Use of the discretized batch-grinding model for the simulation of locked-cycle grinding tests from batch-grinding data is illustrated. The simulated results were compared with actual locked-cycle experimental data for the dry grinding of dolomite. To illustrate the utility of this simulation method, a number of relationships between operating and performance parameters were generated, including relations between batch grinding and the Bond test, the production rate and the circulating load, and the production rate and the product size. A large amount of testing is done each year in the mineral industry for use with grinding unit cperations, and much of the experimental test data is obtained by the tedious locked-cycle testing procedure. In addition to being experimentally complex and requiring large amounts of sample, the very nature of the locked-cycle experiment yields results that cannot readily be interpreted. The attempts that have been made to model comminution systems during the last few years may provide a better and simpler way to obtain and analyze grinding test data; these models involve elements of both fitting to "model equations" and deterministic representations.1-1 The traditional engineering design techniques for comminution systems have relied mainly upon large amounts of industrial operating data that have been correlated with locked-cycle grinding data.',' Bond's empirical mill-design procedure involving locked cycle testing remains the most generally available method to this day. The purpose of the present paper is to show how much more useful data can be obtained from a simple batch-grinding experiment that provides information which subsequently can be utilized with population balance models of the grinding process to simulate locked-cycle behavior. These simulation results can, of course, be used in Bond's general design scheme. However, the combination of batch-grinding tests and the population balance model allows many more conditions to be considered than do single laboratory tests under specific locked-cycle conditions. The use of these simulation techniques for prediction of locked-cycle behavior represents one of the first practical applications for the simulation method in comminution.

Jan 1, 1971

By Joseph C. Mead, Alan Stanley

The MacIntyre Development of the National Lead Co. is located at Tahawus, N. Y. The operations involve the mining and concentrating of a titaniferous iron ore to produce an ilmenite concentrate and a magnetite concentrate. Construction of the MacIntyre plant was commenced during the summer of 1941,when world conditions threatened to cut off the supply of Indian ilmenite. An open pit mining operation was developed and the crushing and milling equipment put in operation in July 1942. A general description of the operation was given in the Adirondack Issue of Mining and Metallurgy for November 1943. The metallurgy of the mill operation was described by Mr. Frank R. Milliken,* Plant Manager, National Lead Company, MacIntyre Development, and presented at the AIME New York Meeting, February 1948. The magnetite concentrate produced in the milling operation was too fine (minus 20 mesh) to be used directly in iron blast furnace operation, and most of the magnetite had to be stockpiled in 1942 and 1943. In 1943, the Defense Plant Corp. built a Greenawalt sintering plant at Tahawus, N. Y., to put the magnetite concentrate in a more suitable form for use in the iron blast furnace. The Greenawalt sintering plant consists of three 10 by 25 ft sintering pans designed to produce 1800 gross tons of sinter per 24 hr. The vacuum to each pan is produced by two Greenawalt fans in series, pulling approximately 30,000 cu ft of air per minute at 50 in. water gauge vacuum. The plant started operation in August 1944. The present plant production averages 25 tons per operating pan hour (approximately 224 lb per operating hour per square foot of grate area) of plus 1 in. sinter. Raw feed to the plant consists of 61 pct magnetite, 4 pct anthracite coal culm, and 35 pct minus 1 in. return fines which are conveyed to a pug mill where the materials are mixed thoroughly and water added to give the mixture 5.5 to 6 pct moisture. The mixed prepared feed is conveyed to two 4 by 10 ft vibrating screens where the minus 1 in. plus 5/8 in. return fines are screened out and discharged into a surge bin for use as a hearth layer. The minus 5/8 in. prepared feed is discharged into another surge bin for use as prepared feed. A charge car, electrically operated, having a capacity of one charge of prepared feed and several charges of hearth layer, lays a thin layer of plus 5/8 in. return fines and 9½ in. depth of prepared feed into the pans. A fluffing roll and a vibrator on the car fluffs and spreads the prepared feed into the pans. An ignition car, electrically operated, ignites the top of the bed with a 30 sec flash burn. The 9½ in. bed sin-ters in approximately 13 min. Dumping the pan, and recharging and igniting the bed requires 2 min. To improve the quality of the ilmenite concentrate produced in the mill and to reduce the amount of titanium dioxide lost in the mill tailings and in the magnetite product, extensive research work and pilot plant operations have been done on grinding the crude ore to minus 65 mesh size (rather than to minus 20 mesh) and concentrating it by a combination of magnetic separation (for magnetite recovery) and flotation (for ilmenite re-covery). These tests have proved successful in increasing ilmenite re-covery and grade. With the development of the ilmenite flotation process to a stage where a full scale flotation plant was in the design stage, the problem arose of handling the 65 mesh magnetite concentrate that would be produced. In order to study and solve the problems of handling and sintering the 65 mesh magnetite in the sinter plant, a pilot sinter plant was secured from John E. Greenawalt. The effect of using 65 mesh magnetite in the sintering operations was then studied on the 2.4 sq ft test pan, operating under conditions as similar to the large plant as could be set up in the laboratory. A series of tests were run in the test pan on present sinter plant feed that had been mixed in the plant pug mill. An average production and an average quality of sinter produced in this series

Jan 1, 1950

By D. W. Fuerstenau, D. A. Sullevan

In the experimental work for this comparison of wet and dry grinding, it was found that the size distributions for wet grinding operations are characterized by a constant value of the distribution modulus, (a; whereas, for dry grinding the distribution modulus decreases slightly with increased size reduction. Wet grindability seems independent of the weight of material in the mill while dry grindability depends on weight. Experimental data and eqmtions for each process are presented for noncleavable isotropic material, for cleavable or material in which there is some form of internal stress, and for energy consumption. The equation ED/EW=BK-? (where ED and Ew are, respectively, energy for wet and dry grinding, k is the size modulus, and B and ? are constants) was found to give the ratio of energy. Dry grinding is generally considered to require more energy than wet grinding. Taggart' states that power consumption in wet grinding ranges from 60 to 90 pct of that required for dry grinding. Bond' states that dry grinding requires one third more power than wet grinding. It is interesting that Rose and Sullivan' note this same difference in power required for wet and dry grinding, but they state that on purely mechanical grounds it is difficult to see any great difference in fundamental principles between wet and dry milling, since dry milling may be regarded as wet milling with a fluid of low viscosity (i.e., air vs water). Yet it must be the viscosity of the fluid that is responsible for the differences between wet and dry grinding. Probably the most comprehensive investigation comparing wet and dry grinding has been that made by Coghill and Devaney. They compared the grind-ability of both chert and dolomite in a 19x36-in. batch ball mill under both wet and dry conditions. In these tests the weight of ore in the mill was varied from 35 to 200 lb and each grinding test was carried out for such period of time that the amount of -200-mesh material in the product was essentially the same for each test with a given material. Cog-hill and DeVaney's data for the grindability of dolomite, which are reproduced in Fig. 1, show that a comparison between wet and dry grinding is com- plex. The comminution of a material by a wet grinding method is characterized by a constant grindability over a wide range of mill loadings. Since grindability is the weight of material comminuted per unit of energy expended, it is equivalent to (w)/ (p)(t), where W is the weight of material being ground, P is the power, and t is the grinding time. In the case of wet grinding, the grindability of dolomite in the Coghill and DeVaney experiments was 0.546 * 0.009 ton per hp-hr when the load of dolomite in the mill ranged from 50 to 200 lb. On the other hand, for dry grinding no such simple relationship was found to exist between power, time, and weight of material in the mill. Dry grinding efficiency appears to depend greatly upon the amount of material within the mill. When the mill contained

Jan 1, 1961

By Robert E. Brand, Strathmore R. B. Cooke

For a study of the bonding that occurs in magnetite pellets during oxidation, cubes of magnetite, hematite, and quartz were prepared, each with one surface polished and nearly optically plane. These faces were fired together under pressure and at various temperatures to determine whether bonding would occur. Sections of surfaces that bonded were examined microscopically. An attempt was made to measure the strength effected by this treatment. EARLIER work1,2 has shown that the strength developed by fired magnetite pellets containing silica or silicates is due to three separate effects. Two of these are associated with the formation of slag in the interior of each pellet at temperatures above 1000°C, where the atmosphere is substantially though temporarily neutral. The third is independent of the silica content and is responsible for the development of a hard shell around the pellet during the initial stages of firing, a condition which permits the pellet to assume the increasing load of the overlying burden and prevents extrusion of the slagged and semi-liquid interior. Pellets made from hematite concentrates do not show these effects unless specific additives are incorporated, and it is these differences in behavior which in large part account for the difficulty in producing satisfactory pellets from such concentrates. Elsewhere it has been stated that "the most important factor responsible for the development of strength in fired magnetite pellets containing no additive is the oxidation of magnetite to hematite and the subsequent recrystallization and grain growth of this constituent."' The object of the investigation reported in the present paper was to study in detail the mechanism described above and to evaluate its contribution to the strength of the fired pellet compared with other possible mechanisms, excluding slag bonding. This was effected by placing together polished faces of two cubes prepared from specific minerals and then heating them under constant pressure at different temperatures both in air and in nitrogen. The materials used were magnetite, hematite, and quartz. The magnetite was coarsely crystalline material from western Montana, the hematite was Itabira ore from Brazil, and the quartz was of optical grade but of unknown origin. A high-grade Lebanon (Pennsylvania) magnetite concentrate was used to prepare briquets for the investigation of bonding strength. Analyses of the iron oxide materials are given in Table I. The specimens for furnace treatment were prepared with a carborundum cut-off wheel, which cut the minerals into cubes approximating % in. along each edge. Care was taken to make the sides as parallel as possible. The edges of the faces to be polished were beveled, and the cubes were mounted in brass rings with sealing wax, this being essential to prevent the specimens from rocking during polishing. To prevent excessive friction, the wax was recessed so that only the brass ring and the face of each specimen would be in contact with the polishing lap. Each specimen was then ground with No. 600 carborundum to produce a level surface before polishing. The Graton-Vanderwilt process was not used for polishing because of the long time required by this method, but several other methods were tried. Cloth laps using levigated alumina or chromium oxide did not produce the desired polish. Satisfactory results were obtained by means of a pitch lap: which was used for all the mineral cubes, with levigated alumina as the polishing medium. The lap was operated in a machine similar to those used in the preparation of single optical surfaces. It was run at low speeds and moderate pressures, and polishing was stopped occasionally to prevent overheating of the lap. A good polish with flat surfaces could be obtained in about 5 min. After polishing the sealing wax was removed either in hot water or in a solvent such as turpentine.

Jan 1, 1955

By D. S. Flett, D. W. West

The solvent extraction of chromium 111 has been studied for the system Cr 111, H,SO., H,O/RNH/RNH., xylene, where the primary amine used was Primene JMT. Rate studies have shown that extremely long equilibrium times are required, ranging from 1 hr at 80°C to 20 days at room temperature. Heating the solution prior to extraction increases the rate of extraction. The variation in the amount of Cr 111 extracted is an inverse function of the acidity of the aqueous phase. Thus, the slow rates of extraction appear to be connected with the hydrolysis of the Cr I11 species. Extraction isotherms for the extraction of Cr 111 have been obtained for two sets of experimental conditions, namely at 60°C and for a heat-treated solution cooled to room temperature. The separation of Fe 111 from Cr 111 and Cr 111 from Cu 11 in sulfate solution by extraction with Primene JMT has been studied and shown to be feasible. A survey of the literature relating to the solvent extraction of chromium showed that, although many systems exist for extraction of Cr VI, only a very few reagents have been found to extract Cr 111. The extraction of Cr III by di-(2-ethyl hexyl) phosphoric acid has been reported by Kimura.' A straight-line dependence of slope —2 was observed between log D,, and the log mineral acid concentration at constant extractant concentration. Since the slope of this plot reflects the charge on the ion extracted, it must be concluded that a hydrolyzed species of Cr III is being extracted. Carboxylic acids generally do not form extractable complexes with Cr III but di-isopropyl salicylic acie does extract Cr 111. Simple acid backwashing of the organic phase, however, failed to remove the chromium. Similar difficulty in backwashing was found by Hellwege and Schweitzer8 in the extraction of Cr I11 with acetyl-acetone in chloroform. The extraction of Cr 111 from chloride solutions by alkyl amines has been reported4-' but the maximum amount of extraction achieved in these studies did not exceed 10%: From sulfate solutions, however, Ishimori" has shown that appreciable amounts of Cr I11 were extracted by amines. The amines used were tri-iso-octyl amine, Amberlite LA-1 (a secondary amine, Rohm & Haas) and Primene JMT (primary amine, Rohm & Haas). The efficiency of extraction with regard to amine type was primary>secondary> tertiary. Appreciable extraction of Cr I11 was recorded for Primene JMT as the aqueous phase acidity tended to zero. The major difficulty with Cr I11 in solvent extraction systems stems from the nonlabile nature of the ion in complex formation. This accounts for the slow rate of extraction generally experienced and the difficulty encountered in backwashing the Cr I11 from the organic phase in the case of liquid cation exchangers. Consequently, the possibility of extraction of Cr I11 as a complex anion is attractive since the backwashing problems should be minimized in this way. From published data, it appeared that the extraction of chromium from sulfate solutions of low acidity by primary amines afforded the best chance of success for a useful solvent extraction system for Cr iii This paper presents the results of a study of the extraction of Cr I11 from sulfate solution by Primene JMT and examines the application of such an extraction procedure for the recovery of chromium from liquors containing iron and copper. Experimental Chromium solutions were prepared from chrome alum in sulfuric acid and sodium sulfate so as to maintain a constant concentration of sulfate ion of 1.5 molar. Solutions of Primene JMT were prepared in xylene and the amine equilibrated with sulfuric acid/sodium sul-fate solutions, of the same acidity as the chromium solution, until there was no change in acidity between the initial and final aqueous phases. The solutions of Primene JMT conditioned in this way were then used for the equilibration experiments. Equilibrations at 25°C were carried out in stoppered conical flasks shaken in a thermostat; equilibrations at all other temperatures were carried out in stirred flasks in a thermostat. After equilibration, the phases were separated and analyzed for chromium. In the tests on the rate of extraction, small samples of equal volume of both phases were withdrawn from time to time and the chromium distribution determined. The chromium analyses were carried out either coloi-imetrically using diphenyl carbazide, or volu-metrically using addition of excess standard ferrous ammonium sulfate and back titration of the excess iron with potassium dichromate. The oxidation of Cr 111 to Cr VI in the case of the raffinate solution was effected by boiling with potassium persulfate in the presence of silver nitrate and, for the backwash solution, by boiling with sodium hydroxide and hydrogen peroxide. Results Preliminary experiments indicated that extraction results were effected by the age of the chromium solution, higher distribution coefficients being obtained with solutions which had been allowed to stand for some time. Consequently a stock solution of chrome alum, 10 m moles per 1 Cr I11 in 1.4 M Na,SO,/O.l M &SO,,

Jan 1, 1971

By R. B. Walker, R. Palmer, L. C. Bryner

The purpose of this investigation was to further study some of the factors affecting the air oxidation of sulfide minerals. Data obtained from laboratory studies on the effect of temperature, concentration of iron salts, pH, and other catalysts on the biological and nonbiological oxidation of sulfide minerals are presented. The oxidation of copper sulfide minerals and subsequent leaching of the soluble copper has long been a profitable operation. The chief oxidizing agent is oxygen from the air. This reaction is influenced by a number of different factors such as temperature, light, hydrogen ion concentration, oxidizing agents, and various catalysts. Sullivan1 found that copper sulfide minerals are slowly oxidized by the air and that dilute sulfuric acid and iron (111) increased their solubility. At that time it was not known that bacteria were involved in the oxidation of the sulfide minerals. It has been shown that pyrite, molybdenite, free sulfur, and the copper sulfide minerals found in the waste rock dumps in Bingham Canyon, Utah, are largely oxidized by chemosynthetic autotrophic microorganisms found in the leaching streams.2-6 Similar organisms have been isolated in other mining areas7,8 and from the effluents from coal rnines.9 The following are some of the questions that have arisen regarding the mechanisms of the reactions involved in the leaching process. How dependent is the leaching operation upon the biological reactions? How does the temperature effect the biological and non-biological reactions? How does the concentration of iron (II) and (111) effect the rate of oxidation of copper (II) sulfide? Does the iron act as a catalyst? Does activated carbon act as a catalyst on the oxidation of sulfide minerals? What effect does the hydrogen ion concentration (pH) have on the above reactions? The purpose of this paper is to report the results of laboratory studies on these problems. EXPERIMENTAL The apparatus used in this investigation was an air lift percolator described in our earlier work.' The percolator consisted essentially of an air lift which enabled the leaching solution and microorganisms to pass freely through a suspension of mineral (substrate) in Ottawa sand. The air lift served the additional purpose of maintaining an ample concentration of oxygen and carbon dioxide in the nutrient (leaching) solution. (See Fig. 1.) The percolators were so constructed that they could be immersed in liquid constant temperature baths for temperature studies. The thermostats were controlled to +0.5°C. The reactions were all run in the dark. The percolators were charged with one to three g of mineral, dispersed in 100 g of Ottawa sand and 100 ml of nutrient or leaching solution. The sand was purified to remove traces of iron by digesting with concentrated hydrochloric acid and washng with distilled water until free from chloride. After drying, the sand was ready for use. The compressed air supply was regulated to 5 psi and bubbled through an aerator containing distilled water, which was maintained at the same temperature as the bath used. The air was filtered through cotton plugs to maintain sterile conditions. The tops of the

Jan 1, 1968

By M. H. Childs, K. E. Merklin

Blast furnace feed specifications are being revolutionized because of the large increase in iron pellet production from low-grade ores of the Lake Superior region. Recent laboratory findings aimed at a better understanding of factors influencing the physical quality of pellets produced from taconite concentrates are reviewed. Tests for determining better product quality, the effects of additives, and production factors are discussed. The current revolution in specifications for blast furnace feed is typified to a great extent by the recent tremendous growth in iron ore pellet production from low grade ores of the Lake Superior region. Indeed, pellets are responsible, at least in part, for many of the stringent new requirements that are being placed on the suppliers of furnace raw materials. As the tonnage of pellets going to the steel mills increases, more and more attention is being turned to the detailed characteristics that make this product an ideal blast furnace feed. Control of quality in pellets is similar to that of any other product. A basic understanding of the factors that are encountered in their production, from the crude ore through to the finished product, is required, along with careful process supervision and frequent quality checks. Normally, chemical grade is not a problem. This usually can be controlled within the desired specifications by minor operating variations in the con- centrator. While future changes in requirements may result in the need for self-fluxing pellets1 or other chemical changes, today's quality demands center primarily around the physical characteristics of the pellets. It is in this area that many of the recent laboratory studies have proven helpful to the operating plants. A better understanding of the physical factors involved in pelletizing has provided the basis for a continuing program of quality improvement. Tigerschiold and Ilmoni2 and Cooke with various associates are responsible for the development of much of the basic understanding of the pelletiz-ing process. These earlier findings have served as a most useful background for this work. LABORATORY PROCEDURES Magnetite concentrates produced from eastern Mesabi Range magnetic taconites were used for all of the work reported herein. The sizing and chemical analysis of a random sample are shown in Table t Additives usually are mixed with the concentrates in taconite plants before pelletizing to improve the strength characteristics of the unfired balls. Ben-tonite, soda ash, and coal are in common use. The bentonite used in these studies was of the Wyoming sodium type; the soda ash was a standard commercia1 brand. Anthrafines were used where coal was called for. The procedure for making batches of balls at the Hibbing Laboratory is as follows: In an especially designed mulling equipment 2800 g of dry ore are mixed for 10 min with the desired additives. Water is added and mixing continued for 1 1/2 min. The charge is removed from the muller and put through a mechanical shredder to break down any lumps that may have formed. sixty previously formed seed balls are placed in the balling drum, which is in fact a 6 x 16-in. diam airplane tire turning at 53 Wm. Two thousand grams of the shredded concentrate and additive mix are fed by hand into the tire at a con-

Jan 1, 1961

By Strathmore R. B. Cooke, Marcus Digre

That calcium will activate quartz for flotation with anionic collectors such as soaps has been known for a number of years,12'3 and the method has been applied to the concentration of various iron ores4 5 6 as well as to other materials. Clemmer gives the permissible pH range for the activation of quartz by calcium as between 9.5 and 12.5 with an optimum value of 11.0, and Hertzog7 has determined the calcium abstraction of iron ores and of Ottawa sand, but otherwise little quantitative data are available. The purpose of the present investigation was to determine those conditions under which dilute solutions of calcium chloride will activate quartz for anionic flotation and to determine the amount of calcium ion adsorbed by quartz. Those variables, the influence of which was primary, such as calcium, sodium, and hydrogen ion concentrations, were studied. Other factors which influenced the results are discussed in the description of the experiments. Three different experimental methods were used to determine the activation of the quartz. The first method was to treat a sample of ground, deslimed quartz with solutions of calcium chloride and sodium hydroxide, and to determine the calcium adsorbed by chemical analysis of the solution. Advantages are that a direct value for the amount of calcium adsorbed is obtained, and introduction of a collector which may affect the adsorption is avoided. The difficulty is to determine the extremely small differences in the calcium concentrations before and after the abstraction. The second method was to determine the activation of the quartz by observing the contact angle obtained on pressing an air bubble against the polished surface of a piece of quartz immersed in a solution of calcium chloride, sodium hydroxide, and sodium oleate. This is the well-known "contact angle" method. The third method was to determine those conditions under which an air bubble will pick up quartz particles immersed in a solution of the desired reagents. Disadvantages common to the sec- ond and third methods are that introduction of a collector is necessary to obtain bubble contact with the mineral; very small quantities of contaminant can give erroneous results; and no information is obtainable concerning the amount of calcium adsorbed. Advantages are that tedious analyses, which at best give approximate results, are avoided, and the conditions, especially for the third method, closely simulate those of actual flotation. Although the most consistent results were obtained with the third method, this paper describes the techniques used and the results obtained from each of the three methods. Distilled water and analytical grade reagents were used throughout except where otherwise stated. Between tests all glass articles were cleaned with a chromic-sulphuric acid mixture and sodium hydroxide. Abstraction Method The raw material for the tests was Ottawa sand sized between 14 and 28 mesh. Batches of 500 g of this sand were dry ground for 1 hr in an Abbe porcelain mill with 2700 g of quartz pebbles. The batches were split down to samples weighing approximately 30 g, and the weight of each sample was adjusted to exactly this value.

Jan 1, 1950

By T. P. Chen, F. W. Bowdish

Studies made with a captive bubble apparatus on the sulfidization and collection by amyl xanthate of true chrysocolla specimens have defined the ranges of pH value and sulfide concentration which permit contact between the bubble and the mineral surface. Titanium compounds were the most effective of the materials found to activate the sulfidization of chrysocolla. With titanium activation, the contact angles and the ranges of pH value and sulfide wncentration giving bubble contact were all increased. Chrysocolla ores were concentrated by flotation. Chrysocolla ores occur at many localities in grade and quantity sufficient to make mining and millin feasible, but no satisfactory method of concentratio has been found. Although chrysocolla may be leached with acid, only those ores without acid-consuming gangue may be leached economically. Because of its potential importance, a study of the conditions nece sary for flotation of chrysocolla has been carried ou The literature contains a few references to flotation of chrysocolla. Two methods were developed by the U. S. Bureau of Mines.1,2 The first consisted of a fatty acid soap and a high xanthate as collectors of chrysocolla from a synthetic ore, while the second involved the use of hydrogen sulfide and xanthate. Ludt and DeWitt3 demonstrated the difference in adsorptive powers of chrysocolla and quartz for bas triphenyl methane dyes and suggested the use of butyl, hexyl or octyl-substituted malachite green as collector. Jackel4 emphasized the effects of combin tions of reagents such as Aerofloat 31, pine oil, and Reagents 404 and 425 with sodium sulfide and zinc hydrosulfite as conditioning agents. Although he reported recoveries of 89% from a synthetic ore and 98% from a natural ore containing azurite, malachite, chalcopyrite and chrysocolla, careful application of Jackel's method to chrysocolla from Tyrone, N.M., failed to give a high recovery. MATERIALS AND TECHNIQUE Samples from Inspiration, Ariz., and Tyrone and Magdalena, N. M., were used for experimentation and verified as true chrysocolla by leaching tests, specific gravity tests and X-ray diffraction. Chrysocolla does not dissolve at pH 4, although malachite and azurite do. Chrysocolla is about half as dense as the copper carbonates. X-ray diffraction analyses by the powder camera method confirmed the samples as true chrysocolla. A captive bubble apparatus, which cast an enlarged image of the air bubble and the mineral surface upon a screen, was used to check on the character of the surfaces. The specimens were prepared by grinding a flat surface on a glass plate using fine abrasive; then they were washed and kept in distilled water until they were to be treated with reagents. Before each reagent treatment, the specimen was carefully checked for cleanliness in the captive bubble apparatus. It was assumed that the surface was clean if, after fine grinding and washing of the specimen, the bubble would not stick. Specimens were handled with glass forceps, and precautions were taken to avoid contamination of the mineral surfaces. Contact angle measurements were carefully made several times on each treated specimen to obtain reliable average values. EFFECT OF pH VALUE AND SODIUM SULFIDE CONCENTRATION In each experiment, a specimen with a freshly ground surface was immersed for 10 min in a solution of sodium sulfide, washed and immersed for 15 min in a solution containing 30 mg per 1 of potassium amyl xanthate. The specimen was then washed again in distilled water and tested for contact angle in the captive bubble apparatus while submerged in distilled water. In this series of experiments, the pH of the sulfidizing solution was varied from 3 to 7, and the concentration of sodium sulfide, containing 60% Na2S, was varied from 50 to 650 mg per 1. Many combinations of pH value and sulfide concentration resulted in no contact between the bubble and the surface, but over a limited range of conditions, contact angles varying from 24ºto 52ºwere obtained. The data in Fig. 1 show sulfidization conditions that lead to bubble contact and those that do not. The region of contact is surprisingly small, which may indicate why flotation of chrysocolla involving sulfidization has proven so difficult in practice. Several features of the system are illustrated in Fig. 1. In the region between pH values of 4 and 6 with sodium sulfide concentrations below about 350

Jan 1, 1963

By J. E. Campbell, A. W. Wesner, M. Pobereskin

This paper summarizes the results of an investigation sponsored by The Coates Steel Products Company. The objective was to determine the relative effects of certain chemical and physical properties on the wear resistance of grinding balls when grinding cement. In making this study, a number of types of balls were used in both dry and wet grinding in commercial ball mills. The radioactive-tracer technique was used in separating the test balls from the rest of the mill charge and in sorting the different types of halls that were all tested at the same time for each of the grinding operations. PROCESSING OF GRINDING BALLS In order to determine whether the properties of grinding balls, namely, composition, microstructure and hardness, can be related to the rate of wear, the halls used in this study were selected to represent various combinations of these properties. It was recognized that the mechanism of wear was not the same in the wet mill as in the dry mill. Therefore, the types of balls tested in each of the two environments were not necessarily the same. However, all of the types of balls evaluated in the dry mill were tested simultaneously for one time period, while those evaluated in the wet mill were tested simultaneously for three different time periods. All of the test balls were separated from the charge when the mill was dumped. All of the balls were of 0.940-inch diameter when charged, and all but the regular production balls were finish ground to this size to eliminate the decarburized surface layers. The steels and processing methods used, as well as the micro-structures and hardness levels of the balls, are described as follows: Regular production forged SAE 1095 steel balls were used in both the wet and dry mills. They were forged to a nominal diameter of 0.940 inch and quenched in oil from the forging temperature. Each retained a partially decarburized surface layer from the forging bar stock; the maximum depth of this layer was about 0.030 inch and the decarburization was negligible where the ball was sheared from the bar. The hardness at the surface was 40 to 46 Rock-well C in the partially decarburized area and 57 to 62 Rockwell C in the areas not decarburized. The hardened zone extended to a depth of about one-third of the radius. Below the decarburized layer, the hardened zone consisted primarily of hard as-quenched martensite with varying amounts of retained austenite. This structure changed to fine pearlite at a depth of about one-third of the radius. In order to eliminate the effect of the decarburized layer, a set of production balls of SAE 1095 steel was forged to about a 1-inch diameter, quenched in oil from the forging temperature, and ground on ball-bearing grinding equipment to a 0.940-inch diameter to remove the decarburized layer. The balls' hardened zones were similar to those in the previous set. The hardness of the finished surfaces was 52 to 60 Rockwell C. A slight amount of decarburization apparently remained after the removal of about 0.030 inch from the surface. Balls of this type were tested in the dry mill only. Another set of forged SAE 1095 steel balls of 1-inch diameter was heated to 1500 F and quenched

Jan 1, 1961

By S. C. Sun, T. Chao

The reaction of gaseous sulfur trioxide with different aluminum-bearing minerals for the conversion of their alumina content into water soluble sulfates has been investigated. Theoretical considerations indicate that this reaction is thermodynamically feasible and that preferential sulfatization can be accomplished. The five minerals tested were gibbsite, boehmite, diaspore, kaolinite, and halloysite. In general, over 90% sulfatization could be achieved in 1 hr when using sodium sulfate as an additive at a temperature between 500° and 600°C. It was found that sulfates of the light alkali and alkaline earth metals exhibited some effectiveness in promoting the sulfa-tizing reaction. Selective sulfatization of alumina in a mixture of calcined halloysite and ferric oxide was accomplished. The gas-solid type of sulfatizing reactions have become increasingly important in recent years for the treatment of the oxides of such nonferrous metals as cobalt,' nickel,' manganese,3 zinc, and copper.4-6 Similar reactions were attempted in the present investigation for the recovery of aluminum from the low grade aluminous ores. Unlike the Bayer process in which silica is a serious contaminant, or the acid processes whereby iron usually causes difficulties, the reactions between gaseous sulfur trioxide and various aluminous minerals could hopefully be selective for aluminum. THEORETICAL CONSIDERATIONS Reactions between aluminum-bearing minerals and gaseous sulfur trioxide have been shown respectively in Table I and Fig. 1 to be thermodynamically favorable for temperatures below 1000°K approximately. Calculations of the thermodynamic functions were where ?G, ?H, ?Cp and ?S are changes in free energy, enthalpy, constant-pressure specific heat, and entropy of the system under consideration, respectively; T is the absolute temperature, R the universal gas constant, and K the equilibrium constant. Thermochemical data so far available for these calculations were based on sources given by T. Chao.7 In Table I, the free energy values indicate that reactions between the gaseous sulfur trioxide and

Jan 1, 1968