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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 X. Wu

A quantitative damage evaluation caused by ground motions associated with strong roof caving is presented. The strong roofs are modeled as either cantilevering or bridging beams over continuous elastic foundations subject to an exponentially distributed vertical stress. Major factors influencing roof caveability and strain energy buildup are identified and investigated. A methodology to evaluate the strain energy accumulation caused by overhanging roof is presented. The potential for mine tremors is determined in terms of the local Richter magnitude ML which depends on the seismic energy transformed from the stored strain energy released when the roof fails. A particle velocity model is applied to estimate the damage area associated with roof collapse. A case study from a Utah coal mine shows that estimations of bump severity and damage area agree well with the field observations.

Jan 1, 1995

By P. Jonsén

Understanding mill charge motion is important. In the charge, the centre of gravity is shifted from the rotational centre of the mill system, and its motion is induced by rotation of the mill, while at the same time the charge creates a torque into the mill system. Breakage of ore particles and wear of liners/ball media are closely linked to the motion. To study these phenomena in a physically correct manner, numerical models for different parts of the mill system are needed. Validations of such models are scarce, because of the difficulty to measure in a tumbling mill. Experimental measurements in a lab mill were done for diverse load cases: varying feed materials, mill fillings, mill speeds and pulp liquids. The mill is set up to directly measure the charge-induced torque. The accuracy is good with relative uncertainty smaller than ±2% for relevant load cases. A full three dimensional numerical model of the whole mill is used to predict induced torque. Agreement between predicted and measured torque at steady-state is good. In addition, the model can accurately predict the mill start-up behavior for torque and mill power. This proves that the model is physically correct, and can be used for full-scale mills.

Feb 27, 2013

By John A. L. Campbell

Although there has been an increased interest in the electrokinetic properties of coal (1) research on the subject is still in an early stage of development (2,3,4,5,6,7) To the author's knowledge, no electrokinetic studies have been made on coal lithotypes - particularly anthracite lithotypes. This paper presents the results of an investigation of the double layer properties displayed at the anthracite/water interface (whole coal and lithotypes) as determined by streaming potential methods, It is hoped that this investigation and related future investigations will provide correlations from which the surface properties of any coal can be predicted and then related to surface dependent processes such as froth flotation and flocculation. Another aspect of this research, undertaken to determine the physical and chemical parameters necessary for anthracite lithotype differentiation, was previously published (8). The number of parameters involved, lithologic, organic and inorganic, and the wide divergence of these parameters within an individual coal limits the practical utilization of the results, however, the results are of value and of interest in guiding theoretical developments. Keeping the limitations in mind, the results are analyzed in terms of their application to froth flotation processes and are used to indicate the feasibility of making a lithotype separation by flotation.

Jan 1, 1969

Mine design and selection of ore processing capacity depend upon a correct estimate of the grade-tonnage distribution. The estimate must reflect what can be recovered and processed. The McCoy/Cove operation in central Nevada, USA, is a gold and silver mine. Case studies of recoverable reserve estimation for gold and silver were made using multiple indicator kriging and the log-affine variance correction. From this model, the grade-tonnage distribution was constructed. Optimal pits were generated by utilizing the distribution, and the results were compared with the pit generated by ordinary kriging. Studies indicate that the proposed method for multi-metal recoverable reserve estimation is more accurate than ordinary kriging in terms of both tons and the grade. Using recoverable reserves for mine design will maximize global recovery of mineral resources.

Jan 1, 1997

By T. H. Ueng

Solving ventilation networks of natural air splitting is a classical problem in mine ventilation. A common approach to this problem is to formulate it based on Kirchhoff's voltage and current laws and obtain the solution by an iterative technique known as the Hardy Cross method. In this paper, it is shown that the problem can be formulated and analyzed as an unconstrained optimization (minimization) problem. The computational experience with the method of conjugate gradients is also discussed.

Jan 1, 1984

By Ronald A. Crowson

Detailed stratigraphic studies of the overburden at North Carolina Phosphate Corporation new South Creek Mine have delineated several aquifers and acquicludes within the 30-meter-thick overburden section of late Pleistocene sands and clayey sands. A major aquifer occurs in the 10-meter-thick sequence of shelly sands of the early Pleistocene Croatan Formation. The water-charged Croatan sands thicken into a paleodrainage surface on top of the Pliocene Yorktown Formation, a major aquiclude. A three-phase dewatering program is projected: 1) improve surface drainage around mine perimeter to reduce both flowage into the pit and rainfall recharge; 2) use a wellpoint system to dewater the upper 8 meters; and 3) use recharge wells along the axis of the paleo- drainage to dewater the overburden below 8 meters.

Jan 1, 1983

By John B. Wright

This paper describes laboratory techniques and subsequent results of Bureau of Mines research to produce chlorination-grade feedstock from an abundant, low-grade, domestic, rock ilmenite ore. The research is part of the Bureau-effort to devise technology that may help decrease U.S. dependence on imported raw materials. A rock ilmenite containing about 46 wt pct Ti02 was smelted in an electric arc furnace with coke, woodchips, and Na2C03 to separate most of the iron as pig iron and to form a low-iron, titanium-enriched slag. The slag was ground, pelletized, oxidized, and then sulfated with mixtures of SO2 and air in a continuous 4.5- to 9.1- kg/h apparatus at 1023 to 1223 K. The sulfated slag, containing 60 to 65 pct TiO2, was reground and leached in water at ambient temperature to decrease the combined levels of the troublesome Ca, Mg, and Mn impurities from 7.0 to 1.7 wt pct. At levels of more than 3 wt pct, these impurities contribute to operating and disposal problems, The final product was upgraded to about 84 wt pct TiO2. This material potentially is suitable for chlorination in a fluidized-bed reactor. The most favorable results were achieved with leached slag sulfated at a feedrate of 4.5 kg/h and at temperatures of 1073 to 1123 K.

Jan 1, 1984

By R. C. Bunge

The success of a mineral separation is commonly assessed from the separation curve, a graph of the recoveries achieved versus the corresponding separator settings. Two values characterize the curve, first the "splitpoint" which is the separator setting that produces a recovery of 50%, and second a measure for the separation sharpness. The splitpoint represents the "mean" while the measure for the separation sharpness represents the "scatter". In modelling the performance of crossflow separations the mean particle behavior is calculated from the force equilibrium that determines the particle trajectory which leads to a deterministic model. The scatter is then represented by a stochastic model. This presentation discusses the modeling of a crossflow separation and some implications with separation theory. The model is shown to successfully predict the performance of a magnetogravimetric separator.

Jan 1, 1992

The knowledge of aggregate size under various chemical and hydrodynamic conditions is very important for selective agglomeration, flotation and water clarification processes. Traditionally, aggregate size measurements are conducted using an ex-situ technique or in a diluted suspension. In either case, the measurement occurs in an environment that does not resemble the actual conditions under which the aggregates are formed. In this study, an on-line scanning laser microscope has been used to study the coagula size of micronized coal suspensions as a function of pH, solids concentration and agitation intensity. A maximum aggregate size of 310 µ m was obtained under relatively quiescent conditions and at pH values near the iso-electric point (iep) of the coal. The aggregate size results compare well with the coagulation efficiency measurements obtained under the same pH conditions.

Jan 1, 1994

By W. J. Wuest

The U.S. Bureau of Mines (USBM) has recently developed a rock mass classification system for evaluating the structural competency of bedded roof strata; the Coal Mine Roof Rating, or CMRR. When compiled with the companion Lotus 1-2-3 spreadsheet, which is simple to learn and use, CMRR data are a useful tool with applications to longwall gate road pillar design, safety hazard mapping, and artificial roof support analysis. This paper demonstrates these applications under actual conditions found in current U.S. mining operations.

Jan 1, 1995

By R. Bhaskar

An airborne dust cloud is an ensemble of particles of varied size, density, shape and states of aggregation. An understanding and quantifying of the dispersion and deposition of the dust generated in mines is essential to the development of measures for reducing worker exposure to dust. In coal mining, the major portion of the dust is generated during the cutting of coal using either a shearer or a continuous miner. The major phenomenon affecting ambient concentration is the deposition of dust. Particle deposition may take place due to the effect of gravitational, inertial, electrostatic, and thermal forces, and mechanisms such as Brownian and eddy diffusion, and inertial impaction. The three major mechanisms of deposition in turbulent flow in mines are diffusion (Brownian and eddy), sedimentation and inertial impaction. The mechanisms determining particle deposition are size dependent.

Jan 1, 1990

By M. G. Hester

The inadequacy of traditional geostatistical methods for estimating some precious metals deposits has become apparent in the last few years. New methods or estimation such as Indicator, Probability and Multigaussian Kriging are difficult to understand and apply. Case studies using a new method of variogram calculation proposed by Srivastava show the method valuable for variogram estimation in precious metal deposits. Some simple, sometimes unorthodox. estimation methods that have proved useful in Some situations are presented.

Jan 1, 1990

By Fredy A. Elorrieta Agramonte, Gabriel von Rickenbach, Roberto Luis Fonseca, David Córdova Rojas, Rolando Romero, Rico Brändle

The mine’s economic life is dynamic by nature. Some factors that create this scenario are the need to recognize new exploitation areas, the evolution of the mineral prices and the development of new technologies to access them. The dynamics of mining development, especially in the underground environment, changes significantly as the stages unfold. In many cases the final solution differs greatly with the initial. The frequent presence of new conditions and factors often cause the mine to develop in a quasi-permanent transition stage, which could be characterized by the generation of dynamic micro-seismic events that could cause rockburst damage in some specific areas. Although they may be infrequent, these dynamic events can have an important adverse impact on mine safety and causes serious losses. Also, changes in the production schedule during exploitation could be a big issue if the problem is not effectively addressed. When the transition of the mine to an active micro-seismic environment begins as an unforeseen scenario, it is frequent to use static ground support systems as they are readily available. Traditionally, the support systems used are reinforced shotcrete with a middle layer of Welded Standard Wire Mesh (WSWM) and Chain Link Standard Wire Mesh (CSWM) of reduced gauges, i.e. wires less or equal than 4 mm in diameter. Although, this support system is used to control static instability problems, there is no good reference of its energy absorption capacity under dynamic load conditions, creating uncertainty in the effectiveness of controlling rockburst events. These support systems are analyzed at the present paper, testing the reinforced shotcrete layer with 4 mm diameter of WSWM and CSWN under dynamic load. Additionally, the influence of the use of high tensile strength steel wire in the meshes (HSWM) compared to the previous ones, is analyzed to assess the scope in controlling rockburst events and keeping the safety of underground environment.

Apr 26, 2022

By A. L. de Vegt

Hydrogen sulfide is an expensive chemical used in nickel, zinc and copper mining and metallurgical operations for selective recovery and concentration of metals from leach water streams, acid plant blow down streams, refinery electrolyte bleeds and precious metal plant bleeds. Hydrogen sulfide is either produced on site by catalytic reduction of elemental sulfur or transported to the site as liquefied hydrogen sulfide (H2S) or as a sodium sulfide solution (NaHS). Biologically produced hydrogen sulfide has economically very interesting application possibilities in the mining and metallurgical industry. Converting sulfates from waste streams into usable hydrogen sulfide by biological reduction results in operational cost savings in existing lime treatment systems. H2S can be produced biologically for less than $500/ton S. For the past ten years Paques has been engaged in the development and installation of treatment systems based on biotechnological processes to remove sulfur compounds from water, air and gaseous streams. Paques has designed and installed a groundwater treatment system for the Budelco zinc refinery in the Netherlands to remove sulfate and heavy metals. The hydrogen sulfide produced is used for recovery of metals, mainly zinc, from the ground water stream. The system has been operating since May 1992 treating a flow of 5000 m3/d. At Kennecott's Bingham Canyon Utah copper mine a pilot plant started operation in November 1995 to develop the process for selective recovery of copper from leach water using biologically produced hydrogen sulfide from sulfate containing groundwater. Other pilot plants are operated at DOE's Pittsburgh Research Center and at Cominco Research focusing on selective recovery of metals.

Jan 1, 1998

By M. D. Ross

Complex combinations of technical, economical, safety and environmental factors are causing mining engineers everywhere to review their operations. Naturally, some of the traditional methods still have a large role to play because they are, in certain circumstances still adequate and are well understood. But, as circumstances change, a new approach is required. The combination of technical and economical reasons for introducing Long- wall mining in some seams in the United States is associated with strata control. Other reasons include simplifications and improvement of ventilation, reduced coal dust, more output from each unit and crew, increased percentage of recovery from each area of the coal deposit, and reduction of rock dusting. The innovations which transformed Longwall mining from a cyclic to a continuous mining system were the armored conveyor and power loading machines, but the crucial breakthrough for application to high production was the development of the powered roof support, which at once gave the system the advantages of mastery of strata control, increased rate of advance, significant reduction of manpower, and increased safety.

Jan 1, 1974

By I. M. Arbuthnot

Over the last five years, a large number of small to medium sized carbon-in-pulp gold treatment plants have been built in Australia, most designed to treat between 250,000 -1,500,000 tpa of ore. In many cases, the primary requirement was to get a plant built and operating in a short period of time and at minimal capital cost, relating to the limited capital resources and tight cash flow positions of these relatively small mining companies. Since the inclusion of both pre-leach and tailings thickeners resulted in an obvious and significant capital cost, most plants were built without thickeners, without doing a detailed cost-benefit analysis on their inclusion, However, the increasing use of High Rate Thickener?s (HRT s) in the mineral processing industries resulted in a reassessment, in some cases, because of their considerably lower cost. This reassessment was triggered primarily by the need to conserve water in arid mining areas, where borefields were costly to install and water was limited. It has become apparent, though, since startup and operation of these installations, that the savings in cyanide consumption is also a very significant factor and can be used as a primary justification for the installation of a HRT in many cases.

Jan 1, 1991

By L. C. Lang

Effective and profitable mining begins with good blasting results. While the importance of sufficient energy to break the rock is recognized, more attention should be paid to the time factor in blasting. Fragmentation, looseness and the shape of the muckpile can be improved and energy levels lowered through the use of proper delays. The long delay is also the most important parameter in establishing stable walls along the final limits of the pit. The time factor in blasting is explained through the rock acceleration process and its practical application is demonstrated by case histories.

Jan 1, 1981

By H. Akgun, J. J. K. Daemen

Design of plugs for abandonment of boreholes and shafts may be governed by the bond strength between the plug and host rock. Presented in this paper are the results of push-out tests on two types of expansive cement grout plugs emplaced in rock salt. The average interface shear strengths range from 2 to 12 MPa. Peak shear stresses (assuming an elastic stress distribution along the interface) are up to eight times higher. Dissolution along the interface may be enhanced by clay inclusions in the salt and appears to reduce bond strengths. Application of the results to the design of plugs for larger openings (e.g., shafts, drifts, or boreholes) is discussed.

Jan 1, 1995

By H. Habenicht, E. Urschitz

Three mining methods for full extraction in flat coal seams - i.e., longwall (L W), shortwall (SW), and rib pillar extraction (RPE) - are com- pared with each other in view of the introduction of a mechanized, self-advancing roof support. Features are shown according to which RPE appears most attractive under certain conditions. In this presentation, the mining methods are outlined and discussed. The new support (Alpine Breaker Line Support, ABLS) is described, and its employment is explained. A trial operation in RPE using the ABLS has been conducted successfully in a South African coal mine. The results and improvements are stated.

Jan 1, 1988