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By T. McCallum, B. Wellbrook, T. Ratnaweera

"Pressure oxidation (POX) leaching of copper (Cu) concentrates continues to be explored as the next generation of leaching processes to sustain existing Cu Solvent Extraction/Electrowinning (SX/EW) operations as easily leachable ores become scarcer. The leach liquors produced, particularly from total pressure oxidation, contain elevated concentrations of Cu and acid, making solvent extraction processing more challenging. Addressing this challenge will require innovative processing techniques and continuous improvement to be successful in achieving satisfactory physical and metallurgical performance in the SX plant. The ability to efficiently process these feeds via SX plays an important role in the continued success of POX technology. Characteristics of a concentrate leach pregnant leach solution (PLS) are discussed, along with best practices that should be considered for processing by SX. Findings discussed in this paper will shed light on ways to reduce process risk and bring additional improvement to the industry with minimal change to the SX plant. INTRODUCTION Hydrometallurgical processing of copper concentrates continues to gain interest from the industry. Treatment by smelting and electro refining continues to dominate, accounting for ~80% of global copper production, but is faced with its own challenges. The case for increased hydrometallurgical processing has been discussed as high grade ore deposits are diminishing, capital costs for smelting are high, and the environmental aspects present challenges for sustainable practice. The hydromet copper industry is also turning a corner as easily leachable copper oxide deposits are depleting, and excess SX-EW capacity exists in certain regions. The transition to hydrometallurgical processing of copper concentrates will take a bit of effort as leaching of copper sulfides is difficult and byproduct credits must be considered. Various POX leaching processes have been studied and proposed to the industry as viable practices. Of the many processes, commercial use has been practiced with the Mount Gordon Copper Process, total and partial pressure oxidation by Freeport-McMoRan at their Bagdad and Morenci, Arizona complexes, the Sepon Copper Process in Laos, and the bioleach BIOCOP™ process in Chile. The experience related to the current paper centers around Cu concentrate leaching performed by Freeport McMoRan (known as CLP). The patented process was developed as an alternative to smelting and refining in 1998. The technology applies to both molybdenum and copper concentrates. Commercial CLP research and development began with Phelps Dodge in 1999 to 2001. The first commercial plant was commissioned in 2003 at the Bagdad operation. Initial operating conditions of the Bagdad CLP were under high temperature conditions before converting in 2005 to medium temperature conditions. In 2007-2009, the CLP technology was commissioned in Morenci, operating medium temperature and high temperature lite conditions until shutting down due to the economic downturn. The Morenci CLP restarted operation in April 2015 due to acid production benefits for the Morenci leach operations. Renewed solvent extraction test work initiated well before restart and has been an ongoing focus for further SX plant optimization."

Jan 1, 2018

By Wisecarver D. W.

Today the longwall is getting increasing attention in the United States as an excellent way to mine coal. This is not to say it is the best way in every mine because conditions, capital costs, and markets may well dictate another method. However, operators with working longwalls are often showing preference by phasing out room-and-pillar sections and replacing them with longwall. In the United States there are about 90 longwalls in operating coal mines and the number is increasing. In Europe, 95% of underground coal mining is by longwall compared with only 5% in the United States. In the Western States about 25% of the underground coal mines use longwall mining. There is logic behind the increase in longwalling. Compared to room-and-pillar mining, straight single-pass longwall has important advantages that show up in underground coal mining in any part of the country.

Jan 1, 1978

By Otis M. Clarke

The 34th Forum on the Geology of Industrial Minerals gave members a chance to see the geology of various stone and ore deposits displayed graphically in quarry walls and also provided an opportunity to hear and see first-rate papers and posters on the geology of industrial minerals. Also, some field trips passed through areas of spectacular scenery. As a bonus, participants saw scenic mines, some of which are "pretty as a picture." Watercolor paintings prove it. At the Oklahoma Forum, the writer displayed a number of paintings of mines and other landscapes seen during previous Forum trips. These paintings and this brief article are part of a crusade to prove that some mines and geologic vistas are scenic wonders and belong in our visual heritage. It is a plea to save some of our best geologic exposures for future geologists, tourists, and artists.

Jan 1, 1999

By G. D. Trump, F. L. Stanley

Spirals have become one of the most popular methods for treating fine coal. Unfortunately, spirals tend to produce an undesirably high specific gravity cut point and often misplace significant amounts of coal and/or rock. In the present study, a process-engineering tool known as linear circuit analysis was used to identify problems with traditional spiral circuits and to suggest alternative configurations. These analyses indicate that a modified rougher-cleaner circuit incorporating a middlings recycle stream offers the best option for improving performance while maintaining a reasonable circulating load. In light of this finding, a full-scale test circuit was installed and evaluated at an existing preparation plant. Data obtained from the plant tests indicate that the new spiral circuit can simultaneously reduce the cut point and improve separation efficiency. Preliminary economic analysis show that the new circuit provides an attractive payback on capital investment.

Jan 1, 1999

By Jin Kim, Jerry C. Tien

Water is still the most economical and popular means of controlling dust underground. It is a process that captures dust by bringing water droplets into intimate contact with coal particles. Wetting agents, also known as surfactants, can be added to greatly facilitate such a process since they tend to reduce the interfacial tension between the coal particles (usually hydrophobic) and the liquid. The surfactants used in the study were first chosen on the basis of their HLB (Hydrophile-Lipophile Balance) value, and later tested in order to select the right type and amount of surfactants to give the best performance. Both static and dynamic tests were later conducted using different types of surfactants to examine their respective wetting effectiveness on coal particles. Experimental results show close correlation with the theoretically calculated collection efficiency that is based on a horizontal tunnel and a solid-cone water spray.

Jan 1, 1993

By Randal J. Barnes

Computer based techniques for solving the Ultimate Pit Limit Problem have appeared in mining literature since the early 1960's. The majority of papers pertaining to this subject can be partitioned into one of two categories: discussion of the pros and cons of a particular algorithm, or a case study presenting specific applications of an algorithm. Subsequently, there has a1 so been considerable debate concerning the best techniques between, essentially, two camps: those favoring the moving cones techniques, and those favoring the techniques which yield true optimal pit designs. The purpose of this paper is not to add to this debate, but rather to present a few new ideas (partially a new look at an old idea), which can be easily implemented to noticeably improve any ultimate pit limit approach independent of the particular algorithm used. These ideas fall under the heading of "bounding the optimum" (Barnes and Johnson, 1980).

Jan 1, 1982

By P. R. Wotruba

The Fortitude gold-silver deposit is related to a "wallrock" copper porphyry system developed within Middle Pennsylvanian to Permian Antler Sequence rocks adjacent to a Middle-Tertiary altered granodiorite intrusive stock at Copper Canyon. Gold-silver ores of the Fortitude deposit occur with disseminated and massive sulfide replacement mineralization of skarn-like or calc-silicated limy horizons of the Antler Sequence contact meta-sedimentary rocks. A major, north-trending, steeply westward dipping normal fault was important as a conduit for hydrothermal fluids responsible for the metallization in the Fortitude area. Gold-silver mineralization is best developed near a marble front where retro-grade chloritization and destruction of prograde calc-silicate mineral phases is most prevalent. Fluid-inclusion studies performed on the Copper Canyon system indicate a wide variation in fluid chemistry during several hydrosilicate stages that ranged in temperature from 500°C to 220°C.

Jan 1, 1986

By William F. Stowasser

The United States has traditionally been the world's largest producer and net exporter of phosphate rock and related fertilizer products with the exception of rock exports from Morocco. In recent years, however, U.S. producers have faced increasing competition from operations in foreign countries, leaving the future position of the U.S. phosphate .industry as a leader in the world markets somewhat less certain. This paper will review the U.S. phosphate rock industry in relation to its known demonstrated reserves, its historical and present production and export levels, and its future mining potential. The paper will include an analysis of the availability of phosphate rock in the United States represented by cost-tonnage relationships. The U.S. phosphate rock industry has for decades provided raw materials at reasonable costs to the phosphate fertilizer industry, and the fertilizer products have been made available to both domestic and international markets. Natural soils normally contain low quantities of phosphorus, and cultivation exhausts the available phosphorus through the removal of harvested products. Yields diminish, and it becomes necessary to apply phosphorus to the soil in an amount at least sufficient to compensate for removal. Raising the level of phosphorus also permits nitrogen and potash fertilizer nutrients to be effective and sustain the high crop yields necessary to support increasing food demands. In only a few areas of the world are soils, climate, rainfall, and low-cost phosphate fertilizer available to assure high agricultural yields in successive years. One of these areas is the central United States and south central Canada. By contrast, in the Soviet Union, the weather conditions in farming areas frequently cause crops to either bake or drown in the field. To attempt to counter the effect of weather on production, the U.S.S.R. has increased use of fertilizer and equipment in recent years. The effect of fertilizer on various crops has been documented, and in all cases, improvements in yield and plant development were noted, as was the ability of fertilizer to mitigate the effects of unfavorable climatic conditions. Phosphate resources can be classified as either identified or undiscovered. In the identified classification, we are principally concerned with demonstrated phosphate resources, or those that have been measured or indicated. In the southeastern United States, the phosphate mining areas of Florida and the Carolinas have been explored and drilled extensively in attempts to locate new deposits. The exploration programs indicate that another Florida Bone Valley Formation or North Carolina Pamlico River Deposit is unlikely to be discovered in the southeastern U.S. coastal plain. Phosphate rock deposits of commercial interest in the western United States are located in Montana, western Wyoming, northern Utah, and southeastern Idaho. The deposits are in the Permian Phosphoria Formation with most of the mining activity occurring in Idaho. Although substantial demonstrated resources have been identified, recovery costs and the field's isolation from markets have restricted growth of the industry in the west. The phosphate rock deposits in Tennessee are well known, and it is estimated that about 29 million metric tons1 are potentially recoverable. Undiscovered deposits of brown rock in Tennessee would likely be quite small.

Jan 1, 1985

By Jean-Michel Rendu

In most mining operations, some degree of selectivity is applied. Whether one deals with surface or underground operation, ore zones, and waste zones are defined not only on the basis of the value of the ore contained, but also taking mining constraints into account. A number of computer techniques are presented which can be used to determine the optimum location of the footwall and hanging wall of ore zones intersected by drill holes. Minimum mining width and minimum internal waste width constraints are taken into account, as well as other economic constraints. The system is presented as an integral part of a larger manumatic ore reserve and mine planning system.

Jan 1, 1982

By Juan L. Yarmuch, Hyam Rubinstein

Open-pit mine planning is a key process in the definition of mineral reserves and their efficient removal. Ore in open-pit mines is extracted bymining sequential pits called pushbacks. Pushbacks need to satisfy complex geometrical constraints to ensure the efficient usage of equipment. To date, no algorithms that both account for those constraints and maximizeNPV, are commercially available. Therefore, pushback design is done by engineers relying on their expertise to incorporate practical features intothe output of commercial tools. In this article, we present a new algorithm that produces close to optimal practical pushbacks and we illustrate its performance in real mines.

Jun 25, 2023

By L. H. E. Weyher

Minerals beneficiation commonly requires initial crushing and grinding so that the mineral phase we seek to concentrate is adequately liberated to obtain a product of specified grade at optimum recovery. To some degree this also is true for coal preparation. In the comminution process we liberate not only the product but also the tailings phase. The true objective is, therefor, to eliminate from the comminution circuit whatever particulate matter has reached specified liberation, regardless of whether the particle is product or tailings material. Overgrinding either of the two phases ties up mill capacity and usually also burdens the downstream processing steps by increasing costs or decreasing recovery, or both. In view of this situation, one would expect concentration steps to be commonly incorporated into the cominution circuit. The reasons this is not done may vary, but probably include the desire to utilize large scale unit operations, e.g., large mills, in conjunction with very simple processes. Hydrocyclones have satisfied this desire but may not always be employed for the right reasons. Hydrocyclones are indeed simple. They also are rugged and forgiving. Unfortunately, these virtues are frequently misused. The direct results are inefficient classification and overgrinding. The indirect results are added costs through recovery losses, unnecessary recirculation, and so on,

Jan 1, 1985

By Craig Smith

California is the nation's leading producer of diatomite products. The Lompoc Mining District in Santa Barbara County is the world's largest source of commercial marine diatomite. Grefco Inc. exploration has defined a large embayment of commercial grade freshwater diatomite in Shasta County. Exploitation of this massive deposit could substantially increase California's diatomite production.

Jan 1, 1980

Geostatistics is well known as a tool for ore reserves calculations at the exploration stage. The goal of this paper is to show geoscatistics can also contribute to productivity gains, allow better selection of the mining methods and better control of the grades during production. The first two examples show how geostatistics can provide sophisticated models, using conditional simulations and the use of them for simulations or optimizations of mining processes. A third example will show how to use geostatistics for short term planning in a large polymetallic vein type mine.

Jan 1, 1983

By A. Fuerst

This paper delivers insights into how high availability combines with high efficiency at the Yanacocha Plant, Newmont Mining Corporation, Peru. Advances in motor design and maintenance with integration of smart control support the approach plant maintenance staff is applying to deliver such excellent results for availability and efficiency in Yanacocha and elsewhere. Detailed efficiency and availability data for 2011 and 2012 in Yanacocha are provided. ABB and Yanacocha teams collaborate closely by combining their experience and continuous development of gearless mill drive procedures and maintenance knowledge to deliver the impressive performance being achieved. The secrets for high performance shared by the Yanacocha maintenance operations point to their systematic, practical adherence to modern asset management and maintenance concepts.

Feb 27, 2013

By Kenneth L. Clifford

We have heard a lot during the recession about the mining industry now competing in a world market for sales of such commodities as copper, lead, zinc, molybdenum, iron ore, and phosphate. If the US mining industry is to survive, do we not have to reduce direct operating costs to a minimum? How significant is labor cost compared to consumables, electricity, fuel, maintenance, materials, and transportation? Ore grade is ultimately the dominant factor; but can we offset its dominance by improving productivity? Is the US mining industry perhaps only 25-30% as productive as it might be? What prevents the industry from being more productive? Can we retrofit older plants to be more productive or does it take a new plant with the largest equipment available for each unit operation? Present Productivity A copper industry study a few years ago covering 13 US copper concentrators ranging from about 10-45 kt/d (11,000-50,000 stpd) showed that productivity did not vary with capacity above 18 kt/d (20,000 stpd). Operating productivity varied significantly, however, from about 180-635 t/d (200-700 stpd) per operating hourly worker in most units, but one operation had 900 kt/d (1,000 stpd). Concentrators with more than 545 t/d (600 stpd) per hourly worker were in the 23 kt/d (25,000 stpd) or more category. Maintenance productivity covered about the same range, but most concentrators were in the 270¬635 t/d (300-700 stpd) per hourly maintenance worker. Again, one company stood well above others by exceeding 1 kt/d (1,100 stpd). Total hourly productivity ranged from 125-280 t/d (140-310 stpd) per hourly worker for most concentrators. The company with the most productive operation, however, was above 455 t/d (500 stpd). Salaried worker productivity for most operations was in the 725 t/d to 1.2 kt/d (800-1,300 stpd) per salaried worker range. The best operation was 2.2 kt/d (2,400 stpd). The end result is, of course, tons per day per total personnel. Most operations fell in the 110-220 t/d (120-240 stpd) range with the best operation at 400 t/d (440 stpd). This is the productivity definition that is most directly related to unit costs. Many of these operations have made some productivity progress

Jan 11, 1983

By Victor F. Hollister

Skam deposits considered as porphyry copper deposits have been classified into three basic prospecting models: (I) the skam deposit associated with a mineralized intrusive (Pima-Mission), (2) the deposit associated with a barren igneous stock (Ertsberg), and (3) mineralized rock that seems to have no clear association with apluton (Phoenix). Pima-Mission, Christmas, and many other copper-bearing skarns developed in carbonate sequences near mineralized intrusives form an important part of the deposit. Historically, copper derived from the mineralized intrusion accounts for less than half of the metal produced by most porphyry mines, as most mineralization occurs in the intruded rocks. Skam mineralization is an important part of wall-rock metal deposition, because carbonate sequences are preferentially replaced by sulfide and skarn minerals. Exploration for skam mineralization is an important phase of porphyry copper exploration. Impure carbonate beds can become better host rocks for skam deposits than pure carbonates, and these in particular should be included in porphyry-exploration programs if the impure strata are cut by the stock. Skam deposits can also form near stocks that are barren of sulfide. In this model, the two best ore guides are strata favorable for mineralization (usually impure carbonate beds) and the intrusive contact. Faults can also act as important ore controls. Some skarn deposits form that have no clear genetic connection with a stock, although intrusive rocks are known in the district. In such cases, the best ore guides are strata and district-wide faulting. Dikes and sills are also common and mineralization can locally prefer their contacts even though the igneous rocks are barren. Skam minerals are influenced by the chemical composition of the carbonate. Forsterite, brucite, tremolite, and talc are common in skams developed in dolomitic beds. Skams derived from calcic beds often contain wollastonite, calcite, stilbite, or idocrase. Diopside, however, is a common accompaniment of copper sulfide in skam regardless of the original composition of the carbonates. Tremolite and possibly other amphiboles and chlorite can occur significantly in the zinc-lead portion of skams. Gamets occur in magnetite, copper, and zinc-lead zones of skams, but are not as diagnostic as diopside for the copper zone or tremolite for the zinc-lead zone. A general desciption of skam-deposit mineralogy can be used to construct prospecting models of porphyry copper skams. An abstract of such a description follows.

Jan 1, 1991

By James Stewart

"THAT WHICH WE CANNOT SEE" INTRODUCTION Often it has been stated that the greatest of all fears is fear itself - fear created only by that which we cannot see or by that which we fail to understand. The elimination or minimization of fear can best be accomplished by a thorough understanding of the underlying causes of the fear. As related to the mining industry the omnipresent fear is the dread of acquiring cancer through accumulated exposure to "that which we cannot see" - radioactivity. The understanding of the resulting exposure effects can best be obtained through acquisition of authentic data and the subsequent analysis of the same by a group of highly qualified specialists. I may point out that the nuclear industry and its source, uranium mining, greatly surpass any other industries in the degree of public awareness as to its potential environmental effects. Public out-cry has always influenced major decisions relating to how industries conduct their operations and much of their effort has resulted in many changes. The issues of health and safety in the mining industry have created considerable public concern, but intense concern to those of us directly involved within the industry. Increasing concern regarding the effect of human activity on the quality of the outer environment of air, land, and water has developed within the past two decades. There is now a growing concern about the health and safety of persons at work in their inner environment, of their working-places in industry. Hearings and investigations conducted in Canada by three appointed "Royal Commissions" have all confirmed that this concern is well justified when focused on the health and safety of workers in our mines. REGULATORY CONTROL OF URANIUM MINING OPERATIONS In Canada, the ownership of mineral resources and the control of their extraction, except for uranium, is vested under the "British North America Act" in the provinces. In all provinces the statutory basis for health and safety provisions in the mining industry can be found in the provincial "Mining Acts" with each province assuming the responsibility to produce its own "Mining Act". In particular, all of the Mining Acts in Canada place upon the management of a mine the responsibility to ensure that provincial provisions are met. The "Federal Register" and mining companies parallel this arrangement in the United States. Equivalent legal

Jan 1, 1981

Maptek, an international development software company, released version 3.2 of its Vulcan three-dimensional mining and geological software package. This latest release is expected to give clients better graphics performance and improved overall solutions. In addition, the company also released the NT version of Vulcan for Unix and personal computers. More than 800 code changes have gone into the 3.2 release, the Lakewood, CO-based company said. Many of these changes were in response to requests from Maptek's customers. Others have been motivated internally by the company to achieve better performance in its Envisage graphics interface. The introduction of hot keys and user-defined menus makes Envisage a more flexible interface and a more productive working environment. A new animation option provides new ways to enhance the organization and display of data. The animated data can include point, string, grid, triangulation and photographic-draped data. The shaft-mapping facility is also included in version 3.2. It enables the cylindrical view of the shaft to be unwrapped and displayed on the screen while the data remains in its true world coordinates. Major upgrades to Vulcan include underground blast design and geotechnical modules. The new blast design module allows the user to develop a ring design in any plane defined by the stope layout. Radial and offset holes can be developed. User-friendly editing tools make adjustments to the initial design a simple task. The user defines the plotting and reporting of rings and plot style templates allow flexibility in reporting. New features in the geotechnical toolbox include options to create daylight windows and pitch lines on the stereo nets. Geotechnical databases can be directly updated from the interpretation of the stereo net. Direct annotation of the symbols on the net is now possible and the stereo net display can be customized. Improvements to the documentation and links within the text have made the on-line help easier to navigate. Maptek also released the NT version of Vulcan soft- ware. The comprehensive design of the software pack- age for Unix and personal computers provides a three-dimensional geological modeling and mine planning system within a natural graphics environment. Users who require the power to manipulate extra- large and complex data sets will still be served by the Unix platform. Users who handle small-to-large datasets can now choose between Unix and personal computer platforms. The NT version will appeal primarily to companies that already have a personal computer network. Response to industry feedback was the basis for developing the NT version. "We want to ensure that our clients receive the optimal product for their needs," said Nicolas Simon, vice president of Maptek in North America. Some advantages of the NT version include: ? Most mining professionals are familiar with the personal computer platform. ? All of the familiar personal computer tools are now available to Vulcan users. ? Resource modeling tools that previously required the powerful Unix workstations have been brought into the personal computer environment. This will help mining companies save money. From a binary data point of view, the personal computer NT version is compatible with other platforms, such as SGIIDECIHPISUN, allowing integrated personal computer and Unix environments For more information, contact Cherie LeDoux, Maptek, 165 So. Union Blvd., Suite 777, Lakewood, CO 80228, fax 303-763-4921, e-mail: info@maptek.com, Internet: http:llwww.maptek.com.

Jan 1, 1997

By N. Arbiter

Introduction The theoretical analysis of tumbling mill energetics and performance has largely neglected mill dimensionality, and, in particular, the importance of the length/diameter (L/D) ratio. This is in spite of the fact that practice varies substantially: geographically, for semi-autogenous and autogenous mills, as between North America and South Africa/Scandinavia, and historically, for overflow ball mills, for which the L/D ratio has increased significantly from the earliest small mills to the largest mills currently. The present study is concerned primarily with the influence of the L/D ratio on the design and operation of overflow ball mills, on the occurrence of the overload phenomenon, and on the limits, if any, it may impose on mill capacities because of critical pulp axial velocity limits. It will be shown that the shape factor is of major importance in this area and that its adjustment to the extent that this is practical should remove the diameter limitations previously postulated for this mill type. Dimensionality in mill design Ball mill shape factors in the period prior to 1927 (Taggart, 1927) averaged 1.1/1 for 29 center discharge mills and 1.0/1 for 30 peripheral discharge mills. With the resumption of new plant construction after the 1930s depression, the Morenci concentrator continued the 1/1 ratio with its 3.1 x 3.1 m (10 x 10 ft) mills. The ratio was increased progressively from then on, reaching 1.6 and 1.8/1 for the largest overflow mills currently. As shown recently (Arbiter, 1989), this is in sharp contrast to autogenous and SAG mill shapes, for which the ratio averages 0.4 in North America. On the other hand, South African practice, starting at the turn of the century with autogenous mills having 4/1 ratios, moved toward 1/1 until recently, when a 2.5/ 1 ratio mill was installed. The reasons given for such divergent practice for mill shape factors are in some respects contradictory and generally inconclusive. The most complete discussion from a practical viewpoint (Dor and Bassarear, 1982) is limited to primary SAG and autogenous mills. Considerations of ball mill dimensionality have had a twofold direction. On the one hand, it has been argued that ball mill efficiencies should increase with increasing diameter and that the specific energy for a particular grind should be reduced accordingly. An inadvertent test of this idea at the Bougainville operation (Burns and Erskine, 1983) resulted in drastic underpowering, which led to failure to reach design capacity until additional mills were installed. This can be taken as strong evidence against any increase in efficiency with diameter. In another direction, it has been argued (Arbiter and Harris, 1982, 1983) that there is a limit to ball mill diameters because of the demonstrable limit to axial flow velocities evident in the overload phenomenon, which as a fact is incontrovertible. But that it places a limit on mill diameters overlooks the evidence given below that appropriate variation in the L/D ratio will permit major increases in diameter, limited only by constructional or economic factors. The present study was directed toward quantifying the overload phenomenon through examination of the influence of mill dimensionality variation and mill operating variables on its occurrence. It is shown that varying the operating conditions, specifically the load fraction and the fraction critical speed, can reduce the risk of overload for existing operations; while appropriate decreases in the L/D ratio can minimize the risk in the design of new circuits. Ball mill overload Ball mill overload is a consequence of the approach to a critical velocity with increasing feed rates or circulating loads. Although the effect has been known for over 50 years, there have been no previous attempts to quantify it. The following description of the ball mill as a flow system is the preliminary to a quantitative analysis: 1) In the absence of a ball load, axial flow of pulp through a mill resembles open channel flow, except for disturbances near the shell due to shell/lifter rotation. 2) For a given ball load (Lf), the void fraction available for pulp hold-up (H) for the ascending portion of the load is approximately 0.4 Lf. In the descending portion, it is greater than this and increases with fraction critical speed (Fc) because of load expansion. 3) Increasing the feed rate increases pulp hold-up and progressively fills the available void space. At a critical flow rate, which depends on system geometry, hydraulic head and pulp rheology, void filling reaches its limit; a pool forms rapidly and fills available space outside the ball load and up to the overflow level. Prior to this, pulp discharges mainly along the ascending rim of the overflow. 4) For a small mill (Lo et al., 1990), it has been shown that with increasing feed rates the critical filling is at or near 50% of the mill volume. Power drops rapidly when this level is reached, as required by the torque formula. 5) The transition to overload is associated with the following phenomena: a) The decrease in power draw. b) Damping of mill sound. c) Reduced comminution of coarser feed sizes, probably due to reduced direct impact in the presence of a pool. d) Increased circulating loads, which further intensify the overload, and increased density of cyclone underflows, which can lead to roping. e) Conditions beyond overload are not known because feed rates are not delibelrately increased beyond this point. f) The existence of the phenomenon limits the capacity of a mill with a fixed set of operating conditions and can prevent the balancing of hard ore feed rate decreases by increases in soft ore rates.

Jan 1, 1992

By H. W. Hitzrot

Introduction Mechanical classifiers consist of a settling tank with parallel sides and a sloping bottom equipped with a mechanism which continuously agitates the pulp and removes the settled solids. The functions served by mechanical classifiers include the following: (1) allow the particles larger than the desired size to settle in the tank and provide an overflow product with a minimum of oversize particles, (2) produce an overflow product of sufficiently high solids content to meet the requirements of the subsequent processing steps, (3) agitate the pulp and permit separation of the entrapped undersize particles so they can report in the overflow product, and (4) drain and remove the underflow solids from the settling pool. Various types of mechanisms have been used to agitate the pulp and remove the settled solids, which include: (I) endless belt and chain drags, (2) reciprocating rakes, (3) helical flights mounted on a rotating shaft which extends through the settling pool, (4) helical flights inside of a rotating drum which contains the classifier pool, (5) troughed conveyor belts forming the settling pool in the concave curve of the belt closely following the tail pulley, and (6) low-speed impellers agitating the underflow settled in the apex of cone-type mechanical classifiers.23 The principal use of mechanical classifiers has been in closed¬-circuit wet grinding; however, this application has been widely displaced over the past few decades by the hydrocyclone. The hydrocyclone installation, including its feed pump, offers lower capital costs and requires less floor space. Nominally it produces an overflow prod¬uct of higher solids content, compared with a mechanical classifier, for comparable classification size up to 400µm (35 mesh). However, these advantages of the hydrocyclone are being gained at the sacrifice of some of the advantages of the mechanical classifier. The mechanical classifier requires less power and has lower maintenance costs than the hydrocyclone and its feed pump. The underflow product from the mechanical classifier can have a higher solids content and contain less entrapped undersize particles than the comparable product from the hydrocyclone. Consequently, it is capable of operating at a higher classification efficiency. A reduced circulating load24, 25 results under these conditions in a closed-circuit grinding application. Other applications for mechanical classifiers include desliming, dewatering, and washing operations. Sand-slime separation may be achieved using one or more stages of mechanical classifiers. Typical applications would include washing and deslimining of concrete aggre¬gate or sand; glass sand; abrasives; oyster shells; phosphate rock; iron, nickel, and chromium ores; alumina; zeolites; solar salt; and precipitates from chemical processing. As an example for the latter, Bitzer26 recommended the use of mechanical classifiers for countercur¬rent cementation-in-pulp of copper from the leach liquor. Similarly, mechanical classifiers are used in uranium processing27 for separation of the barren sands from the pregnant leach solution. Depending on the particle size and the effectiveness of the sand¬slime separation, the sand product from mechanical classifiers may be drained to solids content approaching 84%. Thus, mechanical classifiers are frequently used as dewatering equipment for various applications, such as: 1) Make a product having suitable solids content for subsequent transport by belt conveyor. 2) Provide a drained product to reduce cost of drying. 3) Provide a drained, washed, heavy sand product for return to the circuit. 4) Recover fine values from dilute plant effluents from coal, sand, phosphate-rock, and iron-ore washing plants and from steel-mill flue¬dust scrubbers. 5) Drain residual reagents of a preceding process stage prior to conditioning with reagents of different characteristics, e.g., cationic¬anionic flotation of phosphate rock, iron ore, etc. Design and Operating Features The proper selection of mechanical classification equipment re¬quires that the properties of the solids and liquid are adequately de¬fined. The following descriptive information regarding the solids is desired: feed rate, chemical and physical composition, density, tem¬perature, size analysis, and desired separation size. For the liquid used, the following data would be needed: feed rate, density, viscosity, pH, and temperature. Settling Area. The pool area required to permit a particle larger than the separation size to settle depends on the density and shape of the particle and on the density and viscosity of the pulp within the classifier pool. These criteria of the solids and pulp determine the settling rate which, in turn, determines the settling area required. Determination of the settling rate by long-tube batch tests can be used to calculate the settling area. A theoretical approach for estimat¬ing the settling area requirements is discussed in Chapter 1, Classifica¬tion Theory. This procedure can be used when neither test data nor empirical data from similar operations are available. Data on settling-area requirements and settling rates have been compiled by the manufacturers of mechanical classifiers. The approxi¬mate settling area required for a particular separation size can be estimated by using the curve shown in Fig. 65. This curve applies for closed-circuit grinding applications. For dilute velocity classification in open circuit, desliming, the length of the overflow weir, W, multiplied by the weir's mean distance from the feed opening, E, is used as the design criterion, instead of the settling area, in sizing the classifier. The area requirements, W X E, for dilute velocity classification can be expressed in terms of volumetric flow, Q, and settling rate, Rs, as follows: W x E = Q/(18.06 Rs) where W and E are feet, Q is gallons per minute, and Rs is inches per second. Overflow Weir. The hydraulic head, and thus the velocity at which the overflow crests the overflow weir, is one of the factors controlling the particle size of classification. Consequently, the length of the overflow weir must be designed to provide the overflow velocity which allows the particles larger than the desired size of classification to settle in the working area of the pool. Obtaining the necessary

Jan 1, 1985