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By O. Mellgren, A. M. Gaudin, P. L. De Bruyn

In the January 1956 issue: TP 4137B. Adsorption of Ethyl Xanthate on Pyrite. By A. M. Gaudin, P. L. de Bruyn, and Olav Mellgren. P. 65. Since P. L. de Bruyn is now a member of the AIME, the word "associate" should be removed in describing his status. The address of the third author, Olav Mellgren, should read: 74-2 Chingola, N. Rhodesia. P. 68, second paragraph, last line: the word "300-fold" should be replaced by "3000-fold." Also in the January 1956 issue: TP 4158B. Energy Transfer by Impact. By R. J. Charles and P. L. de Bruyn. The authors wish to express their disapproval of the omission of the abstract from the article, since they believe that an abstract fulfills a very important function in every paper.

Jan 1, 1957

By F. F. Aplan, P. H. de Bruyn

The adsorption density of hexyl mercaptan was measured at the gold-solution and the gold-vapor interfaces. This collector is strongly adsorbed at low concentrations, a monomolecular layer being formed at the gold-solution interface at concentrations as low as 2 x 10'6 moles per liter. A maximum contact angle of 83 ± 3° is observed. From a thermodynamic analysis of the experimental data the changes in surface tension at the gold-gas and gold-solution interface are calculated. The data also suggest that the adsorption density of mercaptan at the gold-gas interface exceeds that at the gold-solution interface. In the equilibrium state, a simple flotation system consists of three phases, the solid to be floated, air, and an aqueous solution in stable contact along a line. During the process leading to this final state, a solid-gas interface is created at the expense of a portion of the liquid-gas and solid-liquid interfaces. This change of state is achieved by addition of a collector and, if necessary, suitable modifying agents to the system. By a combined application of the Gibbs Adsorption and the Young Equations, de Bruyn, Overbeek and Schuhmannl were able to show theoretically that the equilibrium distribution of the collector at all three interfaces determines the magnitude of the contact angle. Their analysis led them to conclude that the equilibrium adsorption density of the collector at the solid-gas interface exceeds its density at the solid- liquid interface. This paper succeeded in focussing attention on the neglected role of the solid-gas interface in flotation. In recent years the application of radiotracer, infrared, electrochemical and electro-kinetic techniques to studies of the solid-liquid interface2-5 enhanced our understanding of adsorption processes at this interface. Except for the measurement of contact angles, the solid-gas interface received no attention in fundamental flotation studies. In an attempt to apply the analysis of de Bruyn, Overbeek and Schuhmann to a simple flotation system, an investigation of the adsorption of hexyl mercaptan onto gold was undertaken. Gold was chosen as the solid phase in order to keep the number of dissolved chemical species in the solution phase to a minimum. Hexyl mercaptan (1 - hexanethiol) was selected as collector because its relatively high vapor pressure at room temperature makes it possible to obtain adsorption measurements at the solid-gas interface. This flotation system consists, therefore, of a pure solid phase (gold), an aqueous solution of a weakly dissociated collector (1 -hexanethiol), and a gaseous phase, nitrogen saturated with respect to both water vapor and mercaptan vapor. EXPERIMENTAL MATERIALS AND METHODS Materials: Liquid 1-hexanethiol (C6H13SH) was obtained from Distillation Products Industries and was further purified by double distillation. The purity of the final product was established by boiling point and refractive index measurements. For the adsorption studies from aqueous solution, the radiotracer element, sulfur-35, (half life 87.1 days, 0.169 Mev beta ray) was synthesized into the mercaptan by Tracerlab Inc., Boston, Mass. A close boiling fraction of the synthesized compound was used in the adsorption studies. The specific activity of the radioactive compound was 10 millicuries per gram.

Jan 1, 1963

By H. R. Spedden, A. M. Gaudin, M. P. Corriveau

A preliminary study of silver ion adsorption by sphalerite in ion exchange column with influent electrolyte marked with radiosilver is described. Rapid ion exchange occurs at first followed by slow reaction controlled perhaps by solid state diffusion. One experiment with radiocopper instead of radiosilver is reported briefly. IT has been known for many years that sphalerite is floated after it has been activated by a metallic cation, particularly copper.' The activation of sphalerite by copper is known to be an exchange of an atom of copper for an atom of zinc at the surface of the sphalerite.2,3 It has been estimated from various guesses of the available mineral surface that the depth of the copper-bearing coating is of the order of thickness of one atom. However, the exact mechanism of the activation process is unknown. A number of metals other than copper act in the same general manner."' " In particular, it has been known that silver replaces zinc at the surface of sphalerite. As copper is usually divalent, whereas silver is monovalent, it may well be assumed that there are differences, perhaps major differences, in the detailed scheme of activation by these two cations. Furthermore, since silver has unusual photoelectric properties and may be a contributor to the fluorescence of sphalerite, a study of its activating mechanism might open up some interesting avenues for exploring these other phenomena. The choice of silver as an activator of sphalerite was suggested also by the fact that its radioactive tracer, silver 110, is particularly convenient in that it has a long half-life and is energetic enough to permit counting through a counter window. The experimental technique selected, therefore, included the preparation of solutions of silver nitrate con- taining a proportion of active silver nitrate sufficient to give convenient counting data. In further considering the method to be used, a desirable choice seemed the use of batch experimentation in which a given solution was allowed to act on a certain lot of mineral for a given length of time at a given temperature, this period of mutual reaction being followed by analyses of the separated solid and solution. But an alternative method seemed more attractive. This is the method now so familiar to the users of ion-exchange resins, which was first made famous by the Russian chromatographer, Tswett.6-8 It consists in forming the mineral into a porous column through which the solution is allowed to flow at a constant rate. The usual method of allowing a batch of solid to react with solution is the equivalent of a single plate distillation operation while the chromatographic or ion-exchange column is a multiple-plate method—the multiplicity of the plates being represented by the reciprocal of the rate of flow of solution through the column. It seemed to us that the chromatographic method had such

Jan 1, 1952

By A. M. Gaudin, W. D. Charles

IN flotation lime is used to depress pyrite. For this purpose it is preferred to caustic soda. The low cost of lime and the widespread availability largely account for this preference. However, there is some evidence' that pyrite is more readily depressed by calcium than by sodium ions at the same pH. To measure the difference in adsorption of calcium and sodium, which is thought to lie at the basis of the differences in depressing action, radionuclides of calcium and sodium have been used. The experimental work has included an evaluation of the effect of dissolved oxygen, of the concentration of calcium or sodium, of the concentration of hydrogen ion, and of the anions present. Preliminary work in the investigation entailed preparation of the pyrite, development of a technique for counting pyrite coated with Ca and Na, the radionuclides used, and development of an apparatus to study the adsorption of calcium and sodium on pyrite. The main experimental work included calcium adsorption and sodium adsorption studies detailed below. The following are the most salient observations made during these experiments. Oxygen in solution increases the adsorptive power of pyrite for calcium

Jan 1, 1954

By Claudio Guarnaschelli

Alteration of the electronic distribution of mineral surfaces by light irradiation has been used to modify the adsorption characteristics of flotation reagents. Energy of a few electron volts is suficient to cause the necessary transitions that either enhance or retard the adsorption phenomenon. The adsorption characteristics of potassium ethyl xanthate on galena and marmatite are presented as functions of the energy of illumination, type of semiconductor, and surface oxidation. In dilute solutions, the adsorption differences are readily detected by controlled kinetic measurements. The correlation between the electronic distribution and the adsorption and desorption phenomena at the solid surface is gaining more attention and acceptance. These interrelations have found a practical application in mineral processing since all of the sulfide minerals display semiconducting properties. By varying the ratio of free electron to electron vacancies, the separation of ilmenite from rutilel and zircon from pyrochlore" has been executed successfully. Plaksin et al.' indicate three possible ways of utilizing the semiconductor properties of a mineral. The first one involves the control of the oxidation-reduction potential of the liquid phase, the second one consists of allowing selected additives to diffuse through the mineral surface, while the third one entails a change in the adsorption and chemical activity of minerals by irradiating their surface. The aim of the present work is to consider the effect of irradiation on the adsorption-desorption characteristics of natural minerals. Since the effect of irradiation is to change the electronic distribution at the surface:

Jan 1, 1971

By S. Usui, I. Iwasaki

The adsorption mechanism of dodecylammonium acetate (DAA) on mercury in potassium fluoride solutions at natural, near neutral pH was investigated. Difler-ential capacity combined with electrocapillary measurements on a dropping mercury electrode were found to provide a sensitive and reliable approach. A shift in the electro capillary maximum (or the point-of-zero-charge) towards positive polarization of the surface due to the adsorption of amine indicated a specific affinity of dodecylammonium ion for mercury surfaces. The amine was shown to adsorb competitively with potassium ion on mercury. Through correlation of the adsorption density of the amine, calculated from the thermodynamic analyses of the electrocapillary curves, and of the diflerential capacity curves, the adsorbed state and the orientation of the amine at the interface were considered. The similarities and the differences in the adsorption behavior of the amine on mercury and on such ionic solids as quartz, silver sulfide, and silver iodide were evaluated. With the promise of flotation providing a cheap and large-scale process for the upgrading of low-grade ores and magnetic concentrates, the iron ore industry has made a sustained search over the years for a flotation reagent of general applicability and high selectivity. Having shelves full of commercial reagents already on hand and anticipating more, both the users and the suppliers are in need of an experimental approach sensitive enough to differentiate the individual characteristics in each group of reagents. Therefore, the adsorption behavior of various flotation reagents on mineral surfaces from aqueous solutions has been a subject of much investigation. The current view of adsorption based on the electrical double-layer theory is formulated on the assumption that certain ions in solution act as potential-determining ions. Their concentrations fix the potential drop across the double layer and their adsorptions govern the surface charge.' Two methods used to study the adsorption

Jan 1, 1971

By L. A. Roe

Magnetic separation of iron ores is one of the fastest-growing segments of the minerals beneficiation industry. The tonnage of taconite ores processed annually by magnetic separation will, in a few years, reach 100 million. Magnetic separation occupies an attractive position in the field of ore beneficiation. It is a simple yet effective method, used for some 150 years and steadily growing more important. This type of beneficiation orginated in 1792, when William Ful-larton was issued a British patent covering the separation of iron ore by magnetic attraction. What is magnetism, and why are some materials capable of being magnetized and others not The study involves many unknowns, and even a partial answer to these questions would be a dissertation in solid state physics and perhaps other fields of science. But the last two decades have revealed basic facts as to why certain materials are magnetic and what happens to them when they are magnetized. Some of the newer studies in solid state physics have developed a theory that can be applied to studies of magnetism. It is claimed that magnetism stems fundamentally from the spin of electrons in atoms that tend to go in pairs, spinning in opposite directions. The atom as a whole can act as a magnet only when there is an imbalance of electronic spin. Whenever an atom has an odd number of electrons, therefore, this imbalance exists. This neutralizing effect explains why a piece of material that contains atomic magnets is not necessarily magnetic. The physicists tell us that iron is composed of many small magnetized regions called domains, which consist individually of millions of atoms. The piece of iron becomes magnetized when an external force lines up these domains in the same direction. Fundamental data developed by physicists working in fields far removed from minerals beneficiation are now available for developing better magnetic separators for ore processing. More Data Needed on Magnetic Properties of Minerals: There is less information on magnetic properties of minerals than there is for certain aspects of magnetic separator design. When the data on magnetic properties is available, it is often directed toward the study of powders for tape recorders and other highly scientific endeavors that are of little use to the beneficiation engineer. W. R. Crane's table of tractive forces, published in 1902, is still a guide in the study of magnetic minerals. Bits of other information regarding electrostatic conductivity, dielectric constant values, and permeability of various minerals have also appeared but are of small practical value in magnetic separation of ore. It is extremely important to obtain pure mineral specimens for making investigations and to record results that may be of use in studies of magnetic processes. Reference to the large volume of data developed by those concerned with analysis and design of permanent magnets soon reveals that these people have extensively evaluated the effect of various impurities on magnetic properties. Much of the science of the ferrite industry is based upon spinel-type structures which are varied by substitution of selected elements or even subtraction of elements, leaving vacant sites in the space lattice. Magnetic separators are used to beneficiate a wide variety of industrial minerals. In this application the relatively large volumes of nonmagnetics are usually the commercial products. The amount of mag-

Jan 1, 1959

By Ellis H. Gates

BENEFICIATION of the manganese oxide ores at Three Kids Mine near Henderson, Nev., has evolved over a period of years. Commercial application of the process is on a secure basis, and an effective working hypothesis has been formulated to help solve the problems encountered in flotation. An outstanding difference between the Manganese Inc. operation and other flotation processes is the necessity for high recovery of slime mineral values. Part of the manganese minerals in the Three Kids ore deposit occurs as primary slime or as minerals that are easily slimed by grinding. High recovery of this slime portion is essential to an economic overall manganese recovery. The aim of much of the research at Manganese Inc. was to recover these slimed and finely ground minerals by flotation. Early Agglomeration Processes: Tunbridge' in 1880, followed by Everson2 in 1886 and Cattermole" in 1904, discovered that under certain pulp conditions precious metals and sulfide minerals could be agglomerated if enough soap or oil were mixed into the pulp slurry. These agglomerated particles, Which under ideal conditions ranged from 1/16 to 1/4 in. diam, were recovered by screening, tabling, and hydraulic classification. Even flotation by air occluded in the agglomkrate particles by vigorous stirring was suggested. Cattermole nearly succeeded in getting his process into commercial operation, but it was finally abandoned, presumably because he could nat compete with the low reagent requirements of the new agitation froth process. Since Cattermole's time Christensen in 1921, Borcherdt4 in 1926, and Chapman and Littleford- in 1934 have attempted to apply the principles of agglomeration. In the light of current research at Manganese Inc. on conditioning of pulp at high energy input levels, it is interesting to note the variety of equipment such as tumbling mills, mixing drums, and shakers described by Borcherdt in his patent to produce the agglomerated particles. Ore Characteristics: Manganese in the Three Kids ore occurs as a 'complex mixture of manganese oxides commonly Known as wad. The manganese minerals have Been identified as psilorflelane, hollandite, coronadite, and pyrolusite with minor amdunts of rhodonite and rhodochrbsite. The ppincipal gangue miherals are quaktz, opal, kaolinite, montmoril-lonite, calcite, gypsum, fcelesitite, and barite. Most of the mangandse ore minerals are either porous and sponge-like in structure or soft clay-like masses composed of fine grained mineral particles approaching the size range of colloids. Consequently the ore minerals from the Three Kids mine have relatively large surface areas and high void content per unit weight. The specific area of various manganese ore samples measured at the California Research Corp. laboratories ranged from 32 to 58 sq m per g. The great surface area and high void content explains in part the unusually high reagent consumption required in the treatment of this ore. Current Milling Practice: At Henderson6 he ore is crushed to —1 in. and ground with rod mills in a closed circuit with spiral rake classifiers and cyclones. Sulfur dioxide solution is added to the pulp as it leaves the grinding circuit, immediately followed by addition of a reagent mixture of soap, oil, and wetting agent in the form of an emulsion. The reagents are initially well mixed into the pulp during their passage through centrifugal pumps to the conditioning equipment. All reagents are added prior to conditioning. The intense conditioning or mechanical stirring of the pulp, in which large amounts of power are consumed, is the principal difference between the flowsheet used at Manganese Inc. and a standard flotation circuit. As much as 38 kw-hr of energy per dry short ton of ore in a pulp containing 18 to 23 pct solids is used for conditioning. After conditioning, the pulp passes to a flotation circuit that consists of one roughing stage, a scav-

Jan 1, 1958

By W. F. Stowasser

downdraft traveling grate process to agglomerate pelletized iron ore concentrates has been successfully demonstrated in a pilot plant at Carrollville, Wis. Work there followed several years of development in the Allis-Chalmers Mfg. Co. laboratories, and the pilot plant phase was carried out in cooperation with Arthur G. McKee & Co., consultants and engineers to the iron and steel industry. End result of the process is conversion of iron ore concentrates into a form which can easily be transported and smelted in the blast furnace. Process Description The first of two process steps incorporates the art of balling and prepares the concentrates for burning. The second step consists of burning the green balls on the grate machine to the hardness required for shipping and handling purposes and for reduction in blast furnaces, see Fig. 1. Facilities are provided at the pilot plant to receive carload quantities of concentrate. The concentrates are loaded into a 50-ton bin direct from railroad cars. Because of the variable moisture content of the concentrates after shipment in an open railroad car it is necessary to repulp and refilter the concentrates to maintain a uniform and proper moisture content for the balling operation. Concentrates are conveyed to slurry tanks, and the slurry, at 50 to 60 pct solids, is pumped to a 4x4-ft drum filter. The filter provides feed of uniform moisture to the plant. Magnetite concentrates are normally filtered to produce a cake containing about 10 pct moisture, a necessary requirement for the following balling operation. The filtered concentrate is conveyed to a rotary bin table feeder which acts as a surge bin for the filter cake and delivers a steady flow of concentrates to the balling drum. It is often desirable to make additions to the concentrates as they are fed to the balling drum. These additives, such as bentonite, increase the strength of the finished green pellet and improve ballability of the concentrate. A vibrating feeder supplies additive to the feed belt, and the additive is mixed with the concentrate in the balling drum. The balling drum, shown in Fig. 3, is 8x3-ft diam. An oscillating cutting bar maintains the lining in the drum by trimming off the buildup of excess concentrate as it forms. The drum is operated in closed circuit with a lx4-ft rod-deck vibrating screen. Undersize pellets or seed pellets from the screen are returned to the balling drum until they grow to the desired size. Size of pellets is controlled by the opening in the screen deck. The formation of pellets in the balling drum is affected by many variables. Some of these are: the size distribution of the feed, the particle shape of the concentrate, the feed rate to the drum, the moisture in the concentrate, the speed of rotation of the drum, the slope of the drum, and the type of trimming obtained with the cutting bar. In this process, attempts are made to control the pellet size within the limits of % to 5/8 in. diam. The screened oversize pellets are conveyed under a coal feeder where sufficient powdered coal is added to the belt to produce desired results in the burning process. The top size of the coal successfully used has been 20 mesh, and anthracite was used in the test program. Fig. 4 illustrates the vibrating screen and the coal feeder. The pellets and free coal are conveyed together to the 5x3-ft diam- reroll drum that rolls the coal onto the surface of the pellets. This drum is also equipped with a cutting bar. The prepared pellets, containing bentonite, water, and surface coal, are elevated to the traveling grate, which consists of a continuous strand of 37 pallets. Each pallet, with a grate bar area 2 ft wide by 1 1/2 ft long, has 14-in. high side plates, Fig. 5. Feeding and distribution of the green balls to the grate is handled by a short conveyor which oscillates back and forth across the 2-ft width of the grate. An adjustable vertical plate located several inches in front of the head pulley of the oscillating conveyor controls the height of the bed and levels the moving bed of pellets. This method of feeding prevents segregation of various size pellets as well as fines and produces a uniform, permeable bed. The pallet train moves under the furnace and across four windboxes, located beneath the pallet frames, see Fig. 2. As the green pellets are deposited on the grate, partial drying of the pellets begins over a 2-ft long updraft windbox. The low temperature air reduces the moisture in the pellets in the lower level of the bed and this operation is essential to prevent sagging of the bed during later stages of the Process. The air used for this drying is recuperated from cooling the pellets on the grate, and supplemental heat, required for starting the Process, is obtained from an auxiliary burner. The pellets are then moved by the grate into the furnace and over an 8-ft windbox, designated as the downdraft waste windbox. Products of combustion are exhausted from this windbox to atmosphere. The furnace, shown in Fig. 6, is constructed with three chambers to provide downdraft drying, preheating, and ignition, respectively, to the pellet bed as it passes through. Overall length of the furnace is 5.57 ft; however, the exterior wall ends may be moved to reduce the length and also adjusted to Obtain the bed height desired, The drying, preheating, and ignition sections of the furnace are supplied with medium temperature

Jan 1, 1956

By W. F. Stowasser

Robert E. Hagen (Oliver Iron Mining Div., U. S. Steel Corp.)—Mr. Stowasser and his associates are to be congratulated on their achievements in developing the design of a pilot agglomerating plant, based on information obtained from experiments made in a laboratory sinter pot, and subsequently constructing and successfully operating this plant. In his Transactions article, May 1955 (vol. 201) Mr. Stowasser stated that the output of his heat treating strand was about 2 Lt per sq ft of grate area per day and that the fuel consumption was 0.7 to 0.8 million Btu per Lt of product. If it is assumed that the burning zone extends from the back of the ignition furnace to the beginning of the recuperation windbox, only about 1/3 of the grate area over the windboxes is utilized for burning pellets, and the actual rate at which partially dried pellets are burned is about 6 Lt per sq ft of grate area per day. It may be interesting to compare the above operating data with that obtained at the Extaca plant, where taconite concentrate is treated on a conventional sin- tering machine using limestone or hydrated lime additives with no recuperation of heat. The entire windbox area is used for burning, and a production rate of 2.5 Lt per sq ft of windbox area per day has been attained at a fuel consumption of 1.8 to 2.0 million Btu per Lt of product, the latter being considerably in excess of that reported by Mr. Stowasser, partly because a circulating load of returns amounting to 28 to 30 pct of the machine feed is used and because no attempt is made to recuperate heat from the waste gases. With the high capital and labor costs prevailing today, it seems that future development of the downdraft grate pelletizing process should be directed toward taking advantage of the high burning rate of predried pellets and increasing capacity, perhaps at the expense of fuel economy, by decreasing the proportion of grate area utilized by ancillary functions. For example, eliminating the cooling zone and using this area for burning would increase the proportion of grate area available for burning from 33 to 50 pct of the total windbox area, and a proportionate increase in capacity would result.

Jan 1, 1956

By R. T. Hukki

The purpose of this paper is to present an analysis of the unit operation of grinding and the circulating load, of the unit operation of classification and the circulating load, and of the two superimposed into one operation to clarify the overall trend of events relating the circulating load to the capacity of the closed circuit grinding system, to the specific surface area of the final fine product, and to the energy consumption in the closed circuit grinding process. The capacity of the closed circuit is frequently, if not always, determined by the classifier, not by the mill." With these words a startling idea was expressed in 1934 by Dorr and Anable.1 Today, some 30 years later, grinding in closed circuit is universally accepted. The basic curve presenting the capacity factor versus circulating load by Dorr and Anable has obviously been so convincing that not a single authoritative investigation regarding its validity seems to have been conducted. The only major subject of discussion appears to have been the relative merits of the 'high' and 'low' circulating loads. It seems to be universally agreed that much still remains to be hoped for regarding sharpness of classification. BASIC CURVES FOR CONSTANT MILL LOADING Fig. 1 presents a series of basic curves for constant mill loadings in function of new feed rate and circulating load. The mill loading indicated by each curve is the sum of hourly tonnages of new feed and circulating load. For example, at point 0 where the new feed rate is 100 t/h and the circulating load 300%, or 300 t/h, the mill loading is 400 t/h. In closed circuit grinding, the new feed rate corresponds to the rate of finished classified material produced. Although at this stage the effect of classification is excluded in construction of the basic curves, the trend indicated by these curves is strongly in favor of closed circuit operation under relatively small circulating loads. A study of the curves reveals that, e.g., a mill operating under a load of 200 t/h at point P where the new feed rate is 50 t/h and the circulating load 300% will also operate under the same total load at point Q where the new feed rate is 100 t/h and the circulating load 100%. In other words, a reduction of circulating load from 300% to 100% would lead to doubling of the new feed rate and doubling of the rate of production of finished material, without a change in the total mill loading. The obvious question is, is it also possible by a reduction of the circulating load in a closed industrial grinding circuit to increase the production of finished classified material and, if so, how and to what extent? GRINDING TESTS BY DORR AND ANABLE In a series of open circuit grinding tests reported by Dorr and Anable1 limestone was ground at various feed rates in a 3-ft ball mill. The original data are reproduced in Table I. It can be seen that with the increased feed rate production of -65-mesh material increased and the energy consumption per unit weight of the -65-mesh product decreased in the reverse ratio. The data shown in Table I have been reproduced for our purposes in Table II for a scale corresponding to grinding in a large industrial mill. This reproduction

Jan 1, 1968

By B. G. Zambre, G. A. Parks, J. N. Roco

Automatic controls have long been a fact of technological life. Some understanding of control systems is rapidly becoming essential to the competence of mineral engineers. Students in undergraduate mineral engineering curricula, however, should not be expected to develop proficiency in the design and interpretation of control systems. They should be provided the opportunity to absorb some of the fundamental concepts and cant of such highly specialized peripheral fields as automatic controls. The present paper describes the components of a simple, flexible and relatively inexpensive control system designed primarily for the use of students in independent study projects. The control system is based on a continuous on-stream optical analyzer analogous to commercially available systems using X-Ray fluorescence analysis. The components can be easily adapted to control other separators and separations. As described here, they have been arranged to control the magnetic separation of a magnetite bearing specular hematite from quartz to produce a quartz product of fixed hematite content using a minimum of power. The control system consists of the magnetic separator* with a vibrating feeder, † a sampler, an optical scanner, a differential amplifier, a millivolt recorder ‡ and a motor driven, metered variable DC power supply for the separator electromagnet. An optional magnet current recorder is useful. The system as arranged for closed loop control is illustrated in Fig. 1. The elements of the loop are identified in automatic control jargon in Table I. The same components can be easily adapted for use in other control modes.§.1 * A Carpco Induced Roll High Intensity Magnetic Separator (Carpco Manufacturing, Inc., Jacksonville, Fla.); modified for external magnet power supply. A Jeffrey-Traylor Vibrating Feeder Type 1B (Jeffrey Manufacturing Co., Columbus, 0.). A Beckman Laboratory Potentiometric Recorder (Beckman Instruments, Inc., 619-A) fitted with adjustable limit switches for external circuit control (Beckman Instruments, Inc.. 71980). Terminology is taken from ref. 1. CLOSED LOOP OPERATION The separation is monitored through the optical reflectivity of the tailings stream which is proportional to the hematite content, THA. The difference, A, between the measured analysis (actually a proportional voltage) and a reference analysis (also a proportional voltage) is continuously recorded. The separation is controlled through changes in the electromagnet current, I, supplied from a motor driven variable source. The two primary variables, ? and I, are linked through adjustable limit switches on the

Jan 1, 1965

By C. M. Marquardt

Tramp iron and steel moving on a conveyor belt cause small currents to be generated in a coil situated in a strong magnetic field, which are converted to an alternating current and are amplified. The output voltage from the amplifier fires a thyratron with a relay in its anode circuit, which actuates a howler and simultaneously drops a spot of marker material on the belt. THE problem of tramp iron removal from moving ore belts is a long standing one. When heavy ore streams are carried on a belt, magnetic pulleys and strong surface magnets fail to remove tramp buried in the ore stream. Tramp iron buried in the ore stream cannot be pulled from the bed by surface magnets and is carried past the magnetic pulley by the falling ore stream. With the increased use of detachable bits the problem of the detection and removal of tramp has become more necessary and difficult. Several types of tramp detectors have been developed. It would serve no practical purpose here to review exhaustively the literature on the subject. Methods of tramp detection used are: (1) The magnetic method, wherein the small current generated by the magnetized tramp passing a Coil is used. (2) The unbalanced oscillator method wherein the tank circuit of a stable radio- or audio-frequency oscillator is unbalanced causing a change of frequency or a change in plate current due to changes in eddy currents, hysteresis, or dielectric because of the presence of tramp. (3) The bridge methods wherein the impedance of one leg of an alternating current inductive bridge is changed due to the presence of conducting tramp. Each of these methods of tramp detection has its field of usefulness. The second and third methods are excellent for oxide iron ores, coal, sand and gravel, grain, etc. However, these latter methods are not suitable for use on ores that contain rich conducting sulphides, since a large piece of pyrite or galena will cause the same detector response as a piece of tramp iron. At the Caselton, Nevada, plant of the Combined Metals Reduction Co., sulphide ores are treated; therefore, it was necessary that a tramp detector operating on only the magnetic properties of the tramp iron be used. Tramp iron detectors of this type operate on the principle that if a piece of tramp iron moves near a coil in a high intensity, steady magnetic field, the presence of the moving tramp iron causes the magnetic flux lines through the coil to change. This induces a mall current in the Coil. In series with the Coil is generally placed a relay sensitive to currents of 1 to 3 microamp. When the tramp iron passes near the coil the flux linkages change, causing this very sensitive relay to actuate. Since the contacts of such a relay are very small it must actuate a second relay capable of breaking a larger current

Jan 1, 1951

By C. M. Marquardt

Tramp iron and steel moving on a conveyor belt cause small currents to be generated in a coil situated in a strong magnetic field, which are converted to an alternating current and are amplified. The output voltage from the amplifier fires a thyratron with a relay in its anode circuit, which actuates a howler and simultaneously drops a spot of marker material on the belt. THE problem of tramp iron removal from moving ore belts is a long standing one. When heavy ore streams are carried on a belt, magnetic pulleys and strong surface magnets fail to remove tramp buried in the ore stream. Tramp iron buried in the ore stream cannot be pulled from the bed by surface magnets and is carried past the magnetic pulley by the falling ore stream. With the increased use of detachable bits the problem of the detection and removal of tramp has become more necessary and difficult. Several types of tramp detectors have been developed. It would serve no practical purpose here to review exhaustively the literature on the subject. Methods of tramp detection used are: (1) The magnetic method, wherein the small current generated by the magnetized tramp passing a Coil is used. (2) The unbalanced oscillator method wherein the tank circuit of a stable radio- or audio-frequency oscillator is unbalanced causing a change of frequency or a change in plate current due to changes in eddy currents, hysteresis, or dielectric because of the presence of tramp. (3) The bridge methods wherein the impedance of one leg of an alternating current inductive bridge is changed due to the presence of conducting tramp. Each of these methods of tramp detection has its field of usefulness. The second and third methods are excellent for oxide iron ores, coal, sand and gravel, grain, etc. However, these latter methods are not suitable for use on ores that contain rich conducting sulphides, since a large piece of pyrite or galena will cause the same detector response as a piece of tramp iron. At the Caselton, Nevada, plant of the Combined Metals Reduction Co., sulphide ores are treated; therefore, it was necessary that a tramp detector operating on only the magnetic properties of the tramp iron be used. Tramp iron detectors of this type operate on the principle that if a piece of tramp iron moves near a coil in a high intensity, steady magnetic field, the presence of the moving tramp iron causes the magnetic flux lines through the coil to change. This induces a mall current in the Coil. In series with the Coil is generally placed a relay sensitive to currents of 1 to 3 microamp. When the tramp iron passes near the coil the flux linkages change, causing this very sensitive relay to actuate. Since the contacts of such a relay are very small it must actuate a second relay capable of breaking a larger current

Jan 1, 1951

By H. Rush Spedden, Arvid Thunaes

Pilot plant test work in 1942 and 1943 showed that by a combination of deslim-ing, fine-size classification, and Sullivan deck concentration it is possible to recover heavy minerals such as cassiterite at least as fine as 10 microns in size. This appreciable improvement in gravity concentration practice has been substantiated by several full-sized plants. IN the past, mills treating ores of tin and tungsten by gravity concentration have recovered very little mineral finer than 325 mesh, although some form of slime concentration has been generally attempted by the use of buddies or round tables. This paper describes a series of pilot plant tests made in 1942 and 1943 in which the use of the Sullivan deck was investigated for the recovery of an appreciable amount of the fine values formerly lost. The investigation was initiated by the U. S. Government in an effort to increase the wartime production of tin. Bolivian milling practice at that time included jigging of the coarse sizes, tabling of the intermediate sizes on conventional shaking tables, and the use of buddies or round tables on the finest sizes. Flotation of the pyrite was employed at different stages of the treatment, depending upon the quantity present and the preference of the operator. Usually good recoveries and the bulk of the production were made in the jigging and tabling operations provided that these sections were not overloaded. Buddies and round tables were used for a small additional recovery if it could be shown that they could pay for their high cost of operation. Extensive test work established that by a combination of desliming, classification, and Sullivan deck treatment, cassiterite as fine as 10 microns (1500 mesh) could be recovered when proper conditions were maintained. This test work has been verified by several full-scale installations. The method is generally applicable to the preconcentra-tion of large tonnages of low-grade ores or tailings for the recovery of a small amount of a valuable heavy mineral. The Sullivan deck (or now the Denver Buckman tilting concentrator), which is the essential part of the method, was devised originally for the recovery of tin from tailings of the Sullivan concentrator, Kimberly, B. C. Since the material treated contained only 0.05 pct Sn, it was impractical to attempt to recover cassiterite finer than 500 mesh.*

Jan 1, 1951

By H. Rush Spedden, Arvid Thunaes

Pilot plant test work in 1942 and 1943 showed that by a combination of deslim-ing, fine-size classification, and Sullivan deck concentration it is possible to recover heavy minerals such as cassiterite at least as fine as 10 microns in size. This appreciable improvement in gravity concentration practice has been substantiated by several full-sized plants. IN the past, mills treating ores of tin and tungsten by gravity concentration have recovered very little mineral finer than 325 mesh, although some form of slime concentration has been generally attempted by the use of buddies or round tables. This paper describes a series of pilot plant tests made in 1942 and 1943 in which the use of the Sullivan deck was investigated for the recovery of an appreciable amount of the fine values formerly lost. The investigation was initiated by the U. S. Government in an effort to increase the wartime production of tin. Bolivian milling practice at that time included jigging of the coarse sizes, tabling of the intermediate sizes on conventional shaking tables, and the use of buddies or round tables on the finest sizes. Flotation of the pyrite was employed at different stages of the treatment, depending upon the quantity present and the preference of the operator. Usually good recoveries and the bulk of the production were made in the jigging and tabling operations provided that these sections were not overloaded. Buddies and round tables were used for a small additional recovery if it could be shown that they could pay for their high cost of operation. Extensive test work established that by a combination of desliming, classification, and Sullivan deck treatment, cassiterite as fine as 10 microns (1500 mesh) could be recovered when proper conditions were maintained. This test work has been verified by several full-scale installations. The method is generally applicable to the preconcentra-tion of large tonnages of low-grade ores or tailings for the recovery of a small amount of a valuable heavy mineral. The Sullivan deck (or now the Denver Buckman tilting concentrator), which is the essential part of the method, was devised originally for the recovery of tin from tailings of the Sullivan concentrator, Kimberly, B. C. Since the material treated contained only 0.05 pct Sn, it was impractical to attempt to recover cassiterite finer than 500 mesh.*

Jan 1, 1951

By H. Rush Spedden, Arvid Thunaes

R. R. Knobler and F. E. Albertson—Following the testwork done by Thunaes and Spedden, a Sullivan deck plant was built for the Colquiri mill. This plant started to operate in April 1945 and continues in successful operation. Since May 1950 the Colquiri mill has operated with a completely revamped flowsheet at a rate of 1000 tons per day. The old Sullivan deck section was retained with this new flowsheet. The "slime plant" in Colquiri consists of one 12-cell Denver No. 18-Sp. Sub-A flotation machine, one 16-ft and one 12-ft hydroseparator, six five-deck Sullivan-frames and six shaking tables for cleaning up the Sullivan deck concentrates, as well as 12 Deister plus three Plat-O tables on which the fine sands, which are not treated on the S-decks, are concentrated. Only —325-mesh fines are treated on the Sullivan decks. The combined hydroseparator overflows are fed to the S-decks. From 1946 to 1949, i.e. with the "old" flowsheet and an average mill feed of 800 tons per day an average of 45 tons per day was treated on four Sullivan rougher frames. The two additional frames were used as cleaners. At times, up to 90 and 120 tons of slimes were treated successfully on this Sullivan deck installation. The average operating data for the S-deck section in Colquiri for the year 1949 were as follows: The tests reported in this paper indicated a tin recovery of 42 pct. The actual plant results for 1949 show a 54 pct tin recovery from the Sullivan decks. The total average quantity of tin produced from the S-deck section in 1949 was 15 tons of tin per month. This corresponds to 3 pct of the total mill recovery. The difference between the 6 pct overall recovery-increase expected after the tests in 1943 and the actual plant results given for 1949 is explained as follows. According to the tests, a total of 170 tons was expected to be routed to the Sullivan decks, but because of certain changes in the flowsheet and mill amplifications,

Jan 1, 1951

By H. Rush Spedden, Arvid Thunaes

R. R. Knobler and F. E. Albertson—Following the testwork done by Thunaes and Spedden, a Sullivan deck plant was built for the Colquiri mill. This plant started to operate in April 1945 and continues in successful operation. Since May 1950 the Colquiri mill has operated with a completely revamped flowsheet at a rate of 1000 tons per day. The old Sullivan deck section was retained with this new flowsheet. The "slime plant" in Colquiri consists of one 12-cell Denver No. 18-Sp. Sub-A flotation machine, one 16-ft and one 12-ft hydroseparator, six five-deck Sullivan-frames and six shaking tables for cleaning up the Sullivan deck concentrates, as well as 12 Deister plus three Plat-O tables on which the fine sands, which are not treated on the S-decks, are concentrated. Only —325-mesh fines are treated on the Sullivan decks. The combined hydroseparator overflows are fed to the S-decks. From 1946 to 1949, i.e. with the "old" flowsheet and an average mill feed of 800 tons per day an average of 45 tons per day was treated on four Sullivan rougher frames. The two additional frames were used as cleaners. At times, up to 90 and 120 tons of slimes were treated successfully on this Sullivan deck installation. The average operating data for the S-deck section in Colquiri for the year 1949 were as follows: The tests reported in this paper indicated a tin recovery of 42 pct. The actual plant results for 1949 show a 54 pct tin recovery from the Sullivan decks. The total average quantity of tin produced from the S-deck section in 1949 was 15 tons of tin per month. This corresponds to 3 pct of the total mill recovery. The difference between the 6 pct overall recovery-increase expected after the tests in 1943 and the actual plant results given for 1949 is explained as follows. According to the tests, a total of 170 tons was expected to be routed to the Sullivan decks, but because of certain changes in the flowsheet and mill amplifications,

Jan 1, 1951

By M. E. Wadsworth, A. S. Peck

Infrared spectra disclose that oleic acid will not adsorb on the surface of pure beryl unless the mineral is first activated with HF. The adsorption of oleic acid on HF activated beryl is attributed to the hydrogen bonding of oleic acid monomers to F-surface bridging sites. These sites are formed as a result of the reaction of HF with chemisorbed water or surface hydroxyl. Characteristic infrared spectra of physically adsorbed oleic acid are correlated with flotation at different pH at temperatures of 25°C and 75°C for colloidal and -60-mesh particles. The usual mode of occurrence of the beryllium mineral beryl, Be3A12Si6O18, is in granite pegmatites as pale green, white, or yellow crystals. A dark green gem variety, emerald, may be found in the wall rock of pegmatite veins or in mica schist. The most important beryl deposits in the United States are in the Black Hills District of South Dakota, and several mountain ranges of central and west-central Utah contain potential commercial sources of this mineral. Several papers have been published concerning the flotation of beryl with different conditioning reagents. Kennedy and O'Meara1 reported on the essential use of HF in the flotation of beryl from different ores. The use of calcium hypochlorite to separate HF activated beryl from feldspar after flotation with an amine collector was made by Runke.2 A report on the activation of beryl with different cations using sodium oleate collector was presented by Viswa-nathan and co-workers.3 Flotation of beryl with sulfonate collector and different activating salts was employed by Fuerstenau and Bhappu.4 The role of the surface charge or zeta potential in the flotation of beryl with different collectors and salts was also extensively investigated.5"7 Previous infrared investigations by the authors8 disclosed that the orthosilicate beryllium mineral, phenacite, reacts with oleic acid to form a chemisorbed oleate monolayer on the mineral surface. Thus, chemisorption is the effective mode of adsorption in the phenacite-oleate acid flotation system. In contrast, neither chemisorption or physical adsorption of oleic acid occur on the surface of pure beryl as evidenced in this study. Efforts were subsequently made to determine the mechanism of activation and collector adsorption for HF pretreated beryl. EXPERIMENTAL PROCEDURES Large, white, terminated crystals of beryl were obtained from Madagascar through Ward's Co., Monterey, Calif. The crystals were broken into -60-mesh fragments in a porcelain mortar. Part of this prepared material was ground to a powder with a Fisher Grinder using an agate mortar and pestle. The powder was transferred to a graduated cylinder and mixed with 500 ml of distilled water. After 5 min, the colloidal suspension was separated from the settled particles by decantation and used as a stock for infrared test determinations. Adsorption tests were performed by first transferring 20 ml of colloidal suspension to 50 ml glass centrifuge tubes. The solids were then centrifuged out and the liquid was discarded. Then, the solids were pretreated with one molar HC1 or one molar HF. After 15 min, the solids were again centrifuged out. The mineral solids were washed free of residual acid by repeated centrifuging and water washing operations. Then, 20 ml of distilled water was added to the solids, and the resultant suspension was used for test purposes. In the preparation of mineral for flotation tests, 1.5 g of -60-mesh beryl was deslimed by removing the colloidal fraction by three successive water washes and decantations. Then, the coarse particles were pretreated with either one molar HCl or one molar HF for 15 min. The mineral was next washed free of acid with distilled water, and 20 ml of distilled water was added to the beryl prior to collector conditioning. The conditioning of mineral with collector was performed in 50 ml centrifuge tubes to which either NaOH or HC1 was added for pH modification. One drop (6.0 mg) of U.S.P. reagent grade oleic acid or 2 ml of 0.01 molar sodium oleate was added to the pulp and stirred at about 1500 rpm for 3 min at a

Jan 1, 1968

By M. E. Wadsworth, L. H. Raby, A. S. Peck

Infrared spec troscopy was used to study the adsorption of oleate collector on three varieties of hematite. Each of the minerals was found to react with either oleic acid or a solution of sodium oleate to form a mineral cation-collector anion chemisorbed layer at the solid-liquid interface. The chemisorption reaction displaced surface hydroxyls and was found to be pH dependent. The infrared band characteristic of the antisymmetrical C = 0 structural group of chemisorbed oleate appears between 1520 and 1540 cm-'. Evidence is presented which shows that molecular filming of the mineral surfaces with un-dissociated oleic acid precedes chemisorption. Infrared absorbance is correlated with the flotation behavior of the mineral. Adsorption processes are usually classified into two types which are distinguished and character- ized by the nature of the bond formed between the adsorbate and adsorbent. The process in which molecules chemically bond to solid surfaces is termed activated adsorption or chemisorption. If only weak van der Waals forces bind the molecules to the surface then the process is termed physical adsorption. Chemisorption is of primary interest in flotation systems since selectivity may be obtained if a single mineral is rendered air-avid by a specific collector mineral adsorption reaction. In previous studies infrared spectroscopy was used to define the mechanisms involved in the adsorption of oleic acid and sodium oleate on fluorite, barite, and calcite. Infrared techniques were particularly useful since adsorbed monolayers were identified in situ. Further, the infrared method permitted identification of reaction products in the solid, liquid, and vapor phases of the systems. The chemisorption reactions displaced F- from fluorite, SO4 from barite, and CO3 from calcite. This investigation was pursued in an effort to characterize and interpret the adsorption processes for synthetic hematite, compact red hematite, and specular hematite (crystalline hematite). Adsorption tests were performed using either oleic acid or

Jan 1, 1967