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By Christopher Haycocks, Lawrence P. Johnson, George M. Neall, James M. Townsend

No method of roof control yet devised has proven to be universally acceptable for the wide range of strata conditions experienced in U. S. coal mines. However, a relatively new innovation, the roof truss, has proven itself capable of preventing roof failures which both conventional and resin bolting could not. Presently, the economics of roof trusses are not favorable when compared to bolting due primarily to drilling and installation procedures, but this is expected to change with improved truss and equipment designs. To realize the full potential of this new method of roof control, design criteria are needed which can optimize the effect of the truss for varying strata conditions. A review of the literature reveals that some work has been conducted on the roof truss. Cambrell and Haynes1 (1970) showed that the roof truss is superior to angled roofbolts installed through a header. Sheorey, Verma, and Singh2 (1973) used statics to prove that the optimum angle of inclination should be greater than 45°. However, to date, no systematic attempt has been made to establish design criteria. In this paper, the problem was approached in three ways. First, model studies were performed to determine general trends of truss support as a function of a number of variables. This provided a better understanding of the support mechanism involved. Secondly, beam theory was employed to calculate the actual support capacity of a truss, realizing the many simplifications, assumptions, and shortcomings of this approach to such a complicated problem. Last, field testing will be utilized to test the hypotheses previously developed and provide hard data to further synthesize roof truss design criteria. Field testing has recently commenced and should be complete before mid 1977.

Jan 1, 1979

By Franklin J. Stermole

INTRODUCTION Economic evaluation of an open pit mine is similar to the economic analysis of any major investment project in any industry. It requires (1) taking into account all of the costs and revenues that are estimated to be incurred over the economic life of the project; (2) crediting the project with all the tax savings and tax credits that will be realized over the project life from tax deductions and credits related to various project costs; (3) charging the project with all taxes to be paid on taxable income over the project life; and (4) properly accounting for the time value of money for all project costs, revenues, taxes, tax savings, and tax credits when making the overall project economic evaluation. These considerations must be taken into account for any valid method of economic analysis such as discounted cash flow rate of return (DCFROR) or net present value (NPV) analyses. All economic analyses fall into one of two categories. ( 1 ) comparison of income-producing alternatives, or (2) comparison of service-producing alternatives. The analysis of alternatives in either of these categories can be made by using correctly any of five basic economic evaluation approaches which are. ( 1 ) present value, annual value, (3) future value, (4) rate of return (ROR) , and (5 ) break-even analysis. Of these evaluation techniques, DCFROR and NPV analyses are used most widely for evaluating income-producing alternatives while present worth cost and equivalent annual cost analyses are used most often for the evaluation of service-producing alternatives. In this chapter income- producing alternative analysis is emphasized, so DCF- ROR and NPV are utilized for evaluating the case studies presented.

Jan 1, 1979

By David K. Hylbert

This study investigated roof falls in room-and-pillar drift coal mines in eastern Kentucky and utilized geologic methods for Predicting roof conditions in advance of mining. In the Highsplint mine, the roof was classified in four categories or conditions: Condition A and B (Laminated Shale); Condition C (Rolling Sandstone with Shale in Pockets); and Condition D (Zone of Major Sandstone Roll(s)). Structure contour and isopach maps of the Darby coal bed and associated immediate roof rocks were used to project axial trends of the sandstone roof. Trend mapping in the mine and on outcrop provided more detailed control for roof projections.

Jan 1, 1979

By B. Y. Ekber, A. S. Gorbachev

The development of economically effective and safe methods of mining thick coal seams allow considerable coal reserves to be used as an energy resource all over the world.

Jan 1, 1979

In his paper, Sastry proposes the use of flotation methods other than the "conventional flotation!' process as possible economical methods of removing mineral fines, The practice of Liquid/Solid separation employing microbubbles to assist or induce the flotation of suspended solids is common to both municipal and industrial waste- water treatment; and is also used as a separation step in some processing industries. The pretreatment (if any) of the stream prior to flotation, the source of the microbubbles, and the design of the separation vessel depend on the application. For example: a metal finishing plant would use an elaborate chemical conditioning scheme to precipitate metal hydroxides for removal by a dissolved air flotation unit operating in the effluent recycle mode, while some oily waste can be separated using a dispersed air flotation cell with no chemical conditioning. At the present time, the "conventional flotation" process employed by the mineral processing industry makes use of mechanical agitation and aeration to generate certain sizes and quantities of bubbles and to mix these bubbles with the pulp. Specific types of chemical conditioning are used to obtain the desired results. For the separation of mineral fines, the type of chemical conditioning, bubble generation and separation device needed to obtain optimum results will have to be determined through research. The point to be made is that while the principles of flotation separation can be applied to a wide variety of applications, the specifics of the flotation process may vary with each application. Sastry's paper has dealt with the many areas of research required to fully understand and utilize the flotation process for the beneficiation of mineral fines.

Jan 1, 1979

Electrowinning U.S. 4,066,520 - Recovery of copper from copper sulfide ore. Finely divided ore concentrate is leached with moderately strong sulfurous acid to solubilize copper ions, the leach solution treated with calcium sulfite to reduce the copper ion concentration by forming copper-precipitate solids therein, the solids filtered off, and and the filter cake dewatered to obtain a slurry of copper-precipitate solids amenable to direct electrowinning of copper therefrom using a described electrolysis apparatus which utilizes high current flow with a minimum of mechanical agitation. R. C. Emmett, Jr., J. K. Dicksa. B. C. Wojcik and F. A. Baczek, assigned to Envirotech Corp. Jan. 3, 1978. 14 pp. (204- 108). Filed: 9- 1-76 (7 19,867).

Jan 1, 1979

By Robert M. Abrams

The passage of an ore pulp through the froth flotation process results in the production, after dewatering, of a finely divided concentrate. This intermediate product must be handled to the subsequent process facilities via beit or other types of conveyors, hoppers, bins, chutes and feeders, and the like. Flotation concentrate has some difficult handling properties such as abrasiveness, sluggishness when moist, or fluidity when dry. These characteristics require the judicious selection or design of handling equipment and the careful engineering of transfer chutes. It is to the special problems which are encountered in this in-plant handling procedure that this paper is addressed. The methods which will be discussed have been developed primarily for the handling of the froth flotation concentrates of copper, lead, and taconite iron ores, but may be generalized, with appropriate modifications, to apply to other similar metallic concentrates. The controlled feeding of concentrates from bins and hoppers is covered along with some of the techniques used to transfer concentrates from one belt conveyor to another. Methods are described for unloading belt conveyors into bins or storage piles. The reclamation from storage piles is also treated.

Jan 1, 1979

By Luckie. Peter T., Deurbouck. A. W., Harold L. Lovell, E. R. Palowitch, James K. Kindig

PART 1: DENSE MEDIUM SEPARATION by E. R. PALOWITCH and A. W. DEURBROUK INTRODUCTION During 1974, 43.9% of the 603,406,000 tons (549,099,000 mt) of bituminous coal and lignite produced was cleaned mechanically, of which 82,280,000 tons (74,875,000 mt) was cleaned by dense medium processes.l The percentage of bituminous coal and lignite cleaned mechanically by dense medium processes increased from 7 to 31 % between 1938 and 1974. Dense medium separations include those coal preparation processes which clean raw coal by immersing it in a fluid having a density inter- mediate between clean coal and reject. As there is a general correlation between ash content and specific gravity, it is possible to achieve the required degree of removal of ash-forming impurities from a raw coal by regulating the specific gravity of the separating fluid. Dense medium processes offer the following advantages over other coal cleaning processes: 1) Ability to make sharp separations at any specific gravity within the range normally required even in the presence of high percentages of the feed in the range of 2 0.1 sp gr units. 2) Ability to maintain a separating gravity that can be controlled with ± 0.005 sp gr units.

Jan 1, 1979

Dr. Wadsworth's discussion of leaching is straightforward and certainly points out some of the critical areas in regards to the fundamental aspects of leaching. I will address myself to the more practical aspects of leaching. As a couple of general comments, very fine particles are often dealt with only when necessary. In many processing applications the sands and slimes are separated with the sand portion being treated by common metallurgical techniques and the slime portion being treated essentially by any technology that will recover some of the values. Ore treatment techniques such as leach/precipitate/float (LPF) have been used in the copper industry to take advantage of leachable copper preferentially reporting to the fine fraction. The second advantage, of course, is the improved metallurgy associated with the treatment of the sand fraction. Cases can be made for sand/slime separations in the aluminum and uranium industry where most of the value reports to the slime fraction and the sand fraction is rejected In the agitated leach systems, fine particles are often the most readily attacked and leached. In dump, vat and in-situ leaching, fine particles cause considerable problems. The fines will tend to restrict the solution flow from portions of the leach feed. This either causes a significantly reduced solution flow or, more probably, the solution will channel, by-passing much of the ore. In either case the result is a decreased metal recovery from ore. The fine particles of principal concern are clays of a flat plate-like shape which can very effectively compact, swell and cause a decreased solution flow. Discrete silica grains or sulfide grains due to their shape, even though they may be very fine, will not cause nearly the problems clays and secondary precipitates can cause.

Jan 1, 1979

By Grant Buma

The purpose of this paper is to present arguments and evidence that there are natural processes at work which tend to minimize the impact of solution mining on the environment. The contaminants have two sources. First are elements which were present in the roll-front deposit in insoluble compounds, for example, uranium in uraninite and coffinite, molybdenum, arsenic, iron, cobalt, nickel, etc. in sulfide minerals, vanadium in micas and clays, etc. Second are elements introduced by the leach liquor. Precipitation of reduced compounds, scavenging of heavy elements and absorption by quartz and clays, can all reduce the amount of heavy elements in solution. If a roll can be demonstrated to be actively precipitating uranium and heavy metal elements, then the presence of precipitants (pyrite and carbon) can be assumed. Criteria for recognizing an active roll are described.

Jan 1, 1979

By Richard D. Hagni

Although the ore microscopist routinely examines polished sections to determine the mineralogy and texture of ores, his importance to the solution of problems peripheral to geology is not always fully realized. The techniques which he uses can be applied with little modification to the solution of a wide variety of beneficiation problem. By extending his studies beyond those normally considered to be of interest to the geologist, he can include examinations of polished sections of selected mill products. This is a logical extension, because the minerals and textures in the polished sections of ores are the same as those in millproducts. The mill dressing engineer should be more aware of the types of beneficiation problem that can be solved by the ore microscopist and what the ore microscopist can do for him. A greater degree of cooperation between the ore microscopist and the mill dressing engineer can be of significant mutual benefit. Ore microscopy holds an advantage over chemical analysis, because the interrelations and combinations ofthe mineral constituents of an ore or mill product can be visually determined. This advantage provides solutions to many beneficiation problem. Locking is a type of beneficiation problem that is particularly amenable to ore microscopy. The percentage, type, and distribution of locked particles can be readily determined with the aid of an ore microscope and an estimate made of the extent of additional grinding required for liberation. The extent of intimately locked types, which no reasonable amount of grinding will liberate, can also be evaluated. Mineral coating conditions, free particles of a deleterious mineral, and free particles of a valuable mineral held in flocculated clusters of gangue are other features that can readily be observed. In some instances, visual examination may show that the problem is not one of locking. Photomicrographs illustrate the variety of beneficiation problems for which the ore microscope is applicable. These problems and their resolution demonstrate the importance of the instrument and show the value of close coordination between the mill dressing engineer and the ore microscopist.

Jan 1, 1979

By Donald W. Gentry, John F. Abel

This paper presents some of the results of a rock mechanics instrumentation program designed for, and implemented at. longwall panel 4N at the York Canyon Mine near Raton, New Mexico. The objective of this particular instrumentation program was to determine the response of the rock mass to longwall mining of a thick coal seam utilizing shield-type supports. Measurements were made in the rock along the ground surface, above the coal seam, in the coal seam, and immediately below the coal seam. Longwall test panel 4N had a total developed length of 762 m (2500 ft) and a face width of 168 m (550 ft), [Fig. 1]. The seam mining height in the panel area was approximately 3 m (10 ft). The average depth of overburden above the panel is 107 m (350 ft) although it ranges from 73-134 m (240-440 ft). Roof support at the face was provided by 116 Hemscheidt 320 HSL shields. These shield-type supports weigh 11.4 metric tons each.

Jan 1, 1979

By A. D. Hill, R. S. Schechter, M. P. Walsh, I. H. Silberberg, M. J. Humenick

There are many uranium deposits which are too deep, too poor in quality or too small in extent to be economically strip mined. It now appears that this uranium can be mined using in situ leaching processes. A number of field tests have been, are being or will be conducted to demonstrate the feasibility of In site leaching. Preliminary results are encouraging, but a major question concerns the restoration of the mine site, which is generally contained within an aquifer. It is necessary to ensure that the quality of the water is not adversely affected by the mining operation. One component of the lixiviant is necessarily a cation and, during the course of mining, the ion exchange sites of the minerals in the mine zone become saturated with this cation. Ammonium is the cation now commonly used. Clearly, any restoration procedure or question involving the rate of cation migration with the groundwater all necessarily involve the process of ion exchange coupled with flow in porous media. The equations describing this process are presented in this paper and numerical solutions obtained. Preliminary results showing the effect of isotherm shape, cation exchange capacity, flow and dispersion are presented. These results demonstrate that in many situations cation migration in ground water will be slow.

Jan 1, 1979

By J. W. Leonard, J. C. Anderson, C. T. Holland

PREPARATION CONTROL IN UNDERGROUND MINING Selective Mining According to district Depending upon the particular geographic loca- tion or seam characteristics, modern raw coal preparation practices can be used to obtain improved product quality and uniformity. Nevertheless, certain coal markets may require only marginal quality while others may require very high quality, and these objectives may be sought by commencing the preparation of coal underground at the "working face." Under- ground preparation involves choosing those portions of the seam which can be cleaned to satisfy predetermined market requirements involving ( 1 ) maintenance of uniformity in chemical quality, (2) elimination of removable impurities, (3) control of sizes and size ratios, and (4) control of moisture. The methods employed to attain these objectives vary widely with the (1 ) seam under development, (2) mining equipment available, and (3) changing trends in the consuming markets. When a mining property is opened and development progresses from one section to another, the (1) petrography of the coal in the seam does not remain uniform, (2) character of roof and floor are subject to change, and (3) character and prevalence of mineral impurities are subject to considerable variation. The quality of coal shipped to market from any mine depends upon intelligent foresight, planning, execution, and supervision. Mining becomes a highly technical art. The quality of coal as determined by chemical and petrographic analysis is without question one of the most important considerations in the development and life of any coal-mining property. Variations in such proper- ties as volatile matter, ash, sulfur, Btu content, and petrography are of considerable interest, and to a limited extent so are such properties as phosphorous content, free swelling index, plasticity, hardness, pressure formation during coking, and ash composition. The maximum and minimum specifications of these characteristics as established by competitive market conditions place definite limits on the marketability of a coal.

Jan 1, 1979

By Milton F. Goodrich

The idea of using water-only cyclones to scalp clean coal from the feed to other devices has recently been gaining in popularity. 1,2,6 An indication of this popularity is that water-only cyclone scalping is already employed in a few commercial plants. Water-only cyclones scalp the feed to a heavy media vessel in the Warwick plant,3 to Deister tables in the Rowland plant,4 and to a heavy media cyclone circuit in the Amherst Coal Company's McGregor plant.5 Despite this trend, we found surprisingly little information in the literature on the performance of water-only cyclone scalping. Therefore, we undertook a study of its technical merits, limiting our investigation to a water-only cyclone scalper in conjunction with heavy-media cyclones. We felt this to be a reasonable combination of equipment, since both processes operate well on the same size coal. First, we developed simple mathematical relationships revealing general characteristics of the system. Next, we compared the clean coal yield and quality obtainable by the system on four different coals with that obtainable by heavymedia cyclone circuit alone and water-only cyclones middlings recirculation circuit alone. This was done by using computer program to simulate the operation of the three systems. Finally, we considered the effect that implementation of the system would have on a typical preparation plant circuit. The System The system chosen for study [(shown schematically in Fig. 1)] comprises two stages: a primary stage (PI) consisting of water-only cyclones, and a secondary stage (P2) consisting of heavy media cyclones. The water-only cyclones separate raw coal at a low specific gravity, producing a relatively small amount of high quality clean coal and a refuse containing much good coal. This refuse is then recleaned in the secondary stage, where the sharper separation of heavy media cyclones, operated at a higher specific gravity, reclaims additional clean coal. Finally, the clean coal from both stages is combined as the system's clean coal, while the refuse from the secondary stage becomes the system's refuse output. Theoretical Analysis Two assumptions lie behind the following theoretical analysis. The first assumption is that partition curves can represent not only the performance of each stage but also the performance of the system as a whole. These curves show partition number as a function of specific gravity, where partition number is either the percent or fraction of a raw coal with a given specific gravity reporting to clean coal. The second assumption is that these curves are independent of the raw coal characteristics. At any specific gravity the system's partition number is solely a function of the partition numbers of each stage. For convenience, we use the following variables in describing this relationship: g = specific gravity, PS(g) = the partition number of the system at specific gravity g, P1(g) = the partition number of the primary stage at specific gravity g, P2(g) = the partition number of the secondary stage at a specific gravity g. For this system, we can show algebraically that the relationship is: [Ps (g) = PI (g) + (1- P1(g)) P2 (9)(1)] With this equation the system's partition curve can be determined if the partition curves of each stage are known. By definition, the partition numbers P1(g) and P2(g) must each be less than, or equal to, one. Placing these constraints on Eq. (1), we arrive at the relations: [Ps(g) % P1(g) and Ps(g) % P2(g)(2)] Thus, at any specific gravity the system's partition number is always greater than that of both stages. [Figure 2] shows typical partition curves. Curve I represents the primary stage, which consists of water-only cyclones operating at a separating gravity of 1.3. (The separating gravity or the

Jan 1, 1979

By R. W. Nash, A. E. Schwaneke, V. R. Miller

The Bureau of Mines developed a detector for controlling sorting devices to separate the copper-bearing fragments from the barren portion of Michigan native copper and western prophyry copper ores. A successful sorter could be used to preconcentrate the ore and eliminate costly hoisting or hauling of waste. An induction balance unit was developed and used in sorting tests on crushed sized - 2.5 + 1.3-cm (- 1 + 112- in.) ores from the Kingston (0.9% Cu) and Centennial (1.6% Cu) mines of Upper Michigan. Concentrates of 4.0% and 7.5% Cu, respectively, were obtained at recoveries of 85 and 80% of the copper in the feed. Eighty percent of the sorted ore was rejected as tailing. Sorting larger fragments improved the separation efficiency. Tests on - 10+ 5-cm (- 4 +2-in.) sized Kingston (1.6% Cu) and Centennial (2.1 % Cu) ore samples yielded concentrates of 4.3 and 4.5% Cu, respectively, with corresponding recoveries of 92 and 91 %. A bout 60% of the sorted ore was discarded as tailing. Porphyry copper sulfide ores also responded to electronic sorting but at a lower separation efficiency. Sorting response was affected by the degree of segregation.

Jan 1, 1979

By J. A. Clum, R. W. Heins, T. D. Tiemann

The use of Na-montmorillonite clay hinders (Western Bentonite) in the pelletizing of iron ore concentrates is well established as is the idea that alternative binders must be found. 1-6 This note summarizes work done7 in evaluating a Wisconsin montmorillonite clay binder. The clay occurs in irregualr beds overlying the Prairie du Chien Dolomite and the St. Peters Sandstone near Hersey, St. Croix County. The raw Wisconsin clay contained 65% quartz and 35% clay minerals. The Na20to CaO ratio was 0.234. Suspension viscosity tests of the raw clay gave results only slightly better than water alone, even after treatment with activator materials. Sedimentation treatment reduced the quartz in the decantate to 5% after 240 min (four hours). Twenty minutes of sedimentation produced a decantate with 27.5% quartz which was found as effective as the 5% quartz decantate in pelletizing behavior. The 20 min decantate was used in subsequent tests. The decantate was filtered and dried at 110°(230°F). The minus 200 mesh portion of this was added directly to the magnetite concentrate. Pellets were made in a laboratory sized disc pelletizer from pastes consisting of ore concentrate, balling water (9%) and binder (clay, or clay + resins, or resins). The pastes were premixed in a Waring blender. The binders were: (1) 1% of a commercially used Western Bentonite; (2) 1 % or 3% of the Wisconsin clay (raw, or 20 min decantate); (3) 20 min decantate washed with activator solutions, (to exchange Na for Ca in the clay); (4) decantate with activator added in the balling water; (5) decantate with organic resins; (6) organic resins. Measurements of the 0.46 m (18 in.) Green Drop Number, the Dry Strength and the Fired Strength (fired at 1260°C (2300°F) for 30 min) were made. The dry and fired compressive strengths were determined on an Instron testing machine using a C-load cell and a F-load cell, respectively. The cross-head speed was 0.004 m/sec (0.1 in./min) for both. The results of each measurement were averaged for 15 pellets. A summary of those results is presented in [Table 1]. More detailed results and procedures are presented in another report.7 In addition to the effects on pellet strength due to various binders the effect of varying magnetite concentrate characteristics is noted in [Table l]. There, a few comparative data are shown for Cornwall, PA Magnetite, which has a generally liner particle size than the Black River Falls magnetite which was used for most of the tests.

Jan 1, 1979

By Charles S. Robinson

The Henderson ore body is east of the Continental Divide in the Front Range of the Colorado Rocky Mountains, about 80 km (50 miles) west of Denver. The ore body is being developed for mining by workings that extend more than 1100 m (3500 ft) below the topographic surface. During construction of the mine workings, ground water in fractures in cystalline rock has been encountered. From 1969 to 1973, Charles S. Robinson & Associates, Inc., in cooperation with the engineering and geologic staffs of the Climax Molybdenum Company, a division of Amax Inc., studied the ground water geology of the Henderson mine development area. Detailed geologic record( of the Henderson mine were available as a result of the exploration, evaluation and development of the mine. Records of the flow, chemical composition and temperature of the ground water at and near the surface, from the entire mine, from portions of the mine, and from exploration holes drilled in the mine, were kept as the mine was developed. Where possible, ground water pressures in drill holes and changes in pressure with time, were recorded. The geologic, hydrologic and chemical data were correlated. The following conclusions were drawn from the correlations: The ground water in the Henderson mine occurs in fractures in Precambrian igneous and metasedimentary rock and in a series of Tertiary intrusive rocks. Two zones of ground water were recognized; an active zone, in which the fracture systems are recharged and discharged annually into the surface water system, and a passive zone, in which the fractures were filled with water in the geologic past. Discharge from the passive zone is through mine workings that intersect the fracture systems. Discharge from a fracture or fracture system is initially high but decreases with time. The rate of decrease in flow (or pressure) is exponential. By measuring the change in flow with time, it is possible to calculate the amount of flow at some given period of time in the future. Based on changes in rate of flow and pressure, and the geology, and assuming rates of mining of the ore, it will be possible to predict the volumes of water that will have to be handled during the life of the mine and the area around the mine that will be drained by the mining operation.

Jan 1, 1979

By Jan Visman, S. J. Aresco

INTRODUCTION The accurate sampling of coal, as with most minerals, is a difficult task. Coal is a heterogeneous material made up of different types of coal and varying amounts of mineral matter. The product as mined may contain all the layers of coal and impurities found in the seam as well as portions of the stratas above and below the coal seam. The preparation plant is the device to size, crush, and/or remove impurities so the coal may be shipped as a salable product. Sampling may be needed at many stages in the preparation process. Although this chapter is primarily concerned with sampling in connection with preparation processes, the principles apply to all coal sampling. The main reasons for sampling in connection with coal preparation are: 1) To determine the washability characteristics of the raw coal by means of float and sink tests. These may be used for the design of a preparation plant, prediction of results to be obtained, or monitoring raw coal delivered to the plant. 2) To check the performance of various operating units in the plant, either by float. and sink tests or chemical analyses. 3) To determine the analytical characteristics of the products produced, such as moisture, ash, Btu, sulfur, etc. HISTORY The earliest accepted methods of sampling coal in the United States were given in US Bureau of Mines (USBM) publications by J. A. Holmes,l G. S. Pope,2 and A. C. Fieldner.3 The American Society for Testing and Materials (ASTM) more or less adopted the methods proposed by USBM in their Standard D-21-1916 and it remained a standard for many years. Briefly, this standard called for a 500 lb (227 kg) sample for coals less than 94 in. (19 mm) in size and a 1000 lb (454 kg) sample for all coals over 94 in. (19 mm) in size. No mention was made of increments, but there was an instruction that the sample should spread over the coal

Jan 1, 1979

Dr. Oldshue has provided a useful summary of a voluminous subject and my task will be to select and direct attention to those areas of mixing theory and practice which are germane to mineral processing in general, and to fine particle treatment in particular. Pulp mixing in impeller actuated reactors occurs in mineral processing in several different applications: (1) emulsification of reagents; (2) mixing pulp with 'reagents as in conditioning for subsequent operations (such as flotation, flocculation or agglomeration) ; (3) high intensity agitation, as in (a) disintegrating and dispersing aggregates, and (b) aggregating particles as in shear flocculation or in selective agglomeration; (4) hydrometallurgy; (5) hydrometallurgy with gas phase (e.g., Arbiter process); . (6) flotation in sub-aeration machines. Attention is invited to 3 above in which two opposite effects can be induced by proper choice of chemical treatment and hydrodynamic conditions. Successful processes for treating mineral fines require excellent aggregate dispersion in order to fully accomplish the liberation stage. Some processes depend further upon selective aggregation following dispersion. The term “dispersion” is often used in different senses which for present purposes will be clarified at the bottom of the next page.

Jan 1, 1979