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By Norbert Patzelt, Johann Knecht

INTRODUCTION Grinding mills have been used in the minerals processing industry for over 100 years. Their dimensions have grown continuously during this time. Besides increasing throughput rates of grinding plants due to the depletion of high grade ores, the lower specific in- vestment costs, as well as reduced operating and maintenance requirements are major reasons for this trend. When selecting new plant equipment one must consider that design principles which have proven their reliability on sizes of today's equipment do not automatically warrant a successful operation on the ever larger size of equipment. Modern calculation methods as for instance the Finite Element method already contribute considerably to the safe design of the huge equipment being built today and are a standard tool of the design engineers. More recently, modern computer programs are also being used in order to size the equipment to meet the process requirements. Today, two design principles are on the market - one which supports the weight of such a unit on trunnion bearings through cast conical endwalls and one which is supported through slipper pad bearings arranged at the circumference of the mill shell (Fig.1). The reason for the development of this alternative grinding mill design can be found in the past. During the sixties and seventies the growing sizes of ball mills with high LID ratios caused many mill failures due to cracked endwalls. The accuracy of the calculation methods as well as the quality standards for castings were not developed to a degree required for such kind of heavy equipment. One way to overcome these problems was the increase of the manufacturing quality standards as well as the introduction of the finite element method based on the analysis of the experience available. The biggest grinding mills being built today are large size SAG mills with cast conical endwalls and trunnion bearings (Fig.2). This is due to the fact that mill manufacturers who had come from the conventional ball mill design adopted these principles as well to their SAG mills. These grinding mills perform well without special concern to the operators. Other manufacturers overcame the problems as mentioned above by eliminating completely the heavy castings and trunnion bearings and the problems associated to it (Fig.1). This design was originally applied to ball mills for the mining and other industries. Due to the success of these shell supported ball mills, this design principle was also applied to SAG mills(Fig.3). Despite of the fact that the majority of today's grinding mills are built to the conventional design it is also interesting to have a look at this alternative. Principles which have proven their reliability on sizes of today's equipment do not automatically warrant a successful operation on the ever larger equipment if bigger mill sizes are realized only based on the pantograph principle. With growing grinding mill sizes, the mass and volume flows through the equipment increases rapidly. Thus it is very important not only to concentrate on the safe design of the structural components of the equipment but as well on the process requirements. The influence of the design on important process parameters of dry and wet grinding plants are discussed thereafter. It shall be shown how modern computer programs can assist in the optimization of the design of components in order to fulfil the operational requirements of such large size equipment. PROCESS REQUIREMENTS OF LARGE SIZE GRINDING MILLS Dry Grinding Mills The world's biggest ball mill is a dry grinding ball mill having a diameter of 6.2m and an overall length of 25,5m with a drive power of 11,200 KW or 15,000HP. This grinding mill dries and grinds gold ore at a rate of 500 tons per hour at a moisture content of up to 9,5%. As shown in Fig.4 this mill was built as a shell supported unit. In fact only this design principle allowed to meet the process requirement. This mill could hardly be built with cast conical endwalls due to the constraints of the trunnion bearings limiting the mill inlet. The following case shows how modern computer programs can help to meet the design criteria of the air system of large size dry grinding plants. For dry grinding plants, the gas flow through the SAG mill has to match the drying, as well as the material transportation require-

Jan 1, 1998

By A. Athousaki

The Ecological Footprint is a resource accounting tool used to address underlying sustainability questions. Aggregates consist of natural, artificial or recycled materials. The production of aggregates is the fastest growing activity worldwide since huge quantities are required by the construction industry. This paper presents a parametric investigation of the ecological footprint of transporting aggregates for urban use as a function of tonnage and distance.

Jan 1, 2008

By Jack W. Eloranta, Michael J. T. Orobona, Peter R. VanDelinder

Ore blending for process optimization is becoming an increasing factor for managing cost and productivity. Economic mining and processing a very low value ore in an extremely competitive market results in constant cost-cummng initiatives. The inherent conflict between the lowest possible mining cost and the lowest possible milling cost is reviewed here in the context of blasting for autogenous mills. Iron ore blending focuses on a number of properties-liberation values, weight recovery and concentrate silica being the most important. Other chemical and metallurgical factors are also commonly controlled. Recent research has fueled an emerging interest in blending on physical ore properties as well. Fully autogenous mills rely on the large, competent fraction of crushed feed to act as grinding media. Previous test work indicated that Hibtac mills require specific amounts of 6- to 10-inch ore. High recirculating loads require a steady influx of large rock fragments. Blast designs, therefore, are presently characterized by wide patterns and low powder factors. Current efforts are aimed at improving mill throughput while reducing overall energy costs. Based on research and published case studies and the advent of optical measuring devices, Hibbing Taconite Company has embarked on a blast optimization program to identify optimum mill feed and to design blast fragmentation goals for each geological unit and each mining area. This paper outlines the preliminary findings of an ongoing, broad-based team effort which requires close cooperation of geologists, mine engineers, crushing and milling personnel. Three specific areas of investigation are reviewed: I) Historical relationships between powder factor and mill performance, 2) Blast fragmentation modeling using the Kuz-Ram model, and 3) Drill

Jan 1, 2004

By Dean Treweek, Anthony Harding

"Creating stable, safe, constructible and cost effective support to openings in segmental linings is often an important component of successful delivery of a tunnel project. The many constraints on successful implementation include operational, spatial, geotechnical, structural, and constructability. Failure to address these adequately can lead to significant impacts on cost, schedule, and safety. With reference to past project experiences, this paper provides a guide to the most commonly deployed solutions and their pros and cons to help designers quickly identify the most appropriate solutions for a particular project, or act as a starting point for developing new and/or project specific solutions. It highlights the challenges that designers and contractors must address together for successful support of tunnel openings, and the constraints that arise. It then describes the common structural forms of support, and how each form can mitigate the constraints, allowing an effective opening support system to be implemented. INTRODUCTION While only a small portion of the total excavated volume of a typical tunnel, cross passages are often one of the most complex elements of a bored tunnel to construct. Cross passages in soft ground typically involve ground improvement, groundwater control, and multiple temporary support measures to ensure safe construction in ground that has been overstressed due to the construction of the running tunnels. Permanent lining construction requires careful detailing of waterproofing and reinforcement to support both long term ground and water loads and support the broken out rings of the main lining. While there is not a great deal of literature specific to cross passage construction, the main issues are discussed in some detail by Lawrence and Taylor (2011). However, the temporary support of the running tunnel segmental linings during cross passage construction is not always given the attention it deserves. The authors have been unable to find much literature on the design of such systems, much less the considerations that have led to a particular solution. Lawrence and Taylor only mention two potential methods of support. It may be easy to dismiss the opening support as an inconsequential item, but selection of the wrong support type can have significant impacts on both cost and schedule."

Jan 1, 2016

By R. P. Killmeyer, R. E. Hucko, P. S. Jacobsen

Cleaning of three coals from the eastern US was tested by five selective agglomeration processes and one selective flocculation process. The coals and processes are described, and the results are reported and compared. Two of the processes, which ground the coal to the finest volume mean diameter, met the stringent test program quality specifications.

Jan 1, 1991

By T. R. C. Aston

Extensive undersea coal reserves, identified in the offshore portion of the Sydney coalfield, Nova Scotia, have resulted in the implementation of a multi-year research program to develop site specific guidelines for the undersea longwall mining operations. At present, uncertainty exists regarding the applicability of the current undersea mining guidelines. Preliminary studies have revealed four potential schemes for monitoring seafloor longwall subsidence profiles: seasurface, subsea, geophysics, and direct monitoring. However, further work has indicated that the use of marine geophysical techniques may be the most effective method of identifying and evaluating seafloor subsidence profiles by comparing pre- and post-mining seafloor topographies.

Jan 1, 1990

By William Lyman

GENERAL DESCRIPTION Shrinkage or shrinkage stoping refers to any mining method in which broken ore is temporarily retained in the stope to provide a working platform and/or to offer temporary support to the stope walls during active mining. Since ore "swells" when broken, it is necessary to shrink the muck pile a corresponding amount by draw¬ing some of the broken ore out as the stope is advanced-hence the name. Broken ore retained during stoping is drawn out after the stope has reached its limits. The stope may be left empty or may be filled with waste contemporaneous with, or subsequent to, the final draw. Traditionally the method implies conventional overhand stoping methods with miners working between the muck pile and the stope back, in a space which advances updip with mining and is maintained by balanc¬ing "swell" with "shrink." The shrinkage classification is also applicable to so-called "semishrinkage" methods in open pillar-supported stopes where broken ore is temporarily retained as a working platform but offers no wall support; and to various blasthole shrinkage methods which utilize broken ore temporarily retained in the stope for wall support, but which do not require miners to work from muck pile in the stope. The method is generally applied to steeply dipping veins of strong ore between strong walls. APPLICATION Geometry The geometry of a shrinkable vein is described in terms of dip, width, and regularity along dip. Overall strike and dip dimensions and irregularities along the strike generally impose no restrictions on the method. Dip is ideally 1.2 to 1.5 rad (70 to 90°). As dip falls below 1.2 rad (70°), the shrinkage draw begins to strongly favor the hanging wall side, thus leaving a poor working platform for conventional overhand work. This is particularly true in relatively wide stopes. The sup¬port afforded to the hanging wall also diminishes with decreasing dip, reaching nil as the dip approaches the repose angle of broken ore. Dips below 0.78 to 0.87 rad (45 to 50°) are not generally shrinkable except by open stope "seinishrinkage" methods. Minimum mining width is fixed by working space requirements in the stope-generally about 1 m. Shrink¬age in narrower veins requires that waste rock from one or both walls be broken with the ore and the attendant dilution accepted to achieve the minimum width. Nar¬row stopes are less suitable, encouraging hang-ups and bridging of broken ore, with the attendant problems of erratic draw and incomplete recovery of broken ore. Maximum practical width may be 3 m or less to over 30 m, depending upon the competency of the ore and its ability to stand unsupported across the stope back. This is a vital safety consideration in conventional over¬hand stopes, but is much less of a factor in blasthole shrinkage methods. Very wide veins and massive ore bodies have been mined by transverse vertical shrinkage panels separated by transverse vertical pillars which are either abandoned or recovered later by other methods. Regularity along the dip is a prerequisite of shrink¬age as there must be no serious obstruction to the flow of broken ore downward through the stope to the sill level. Gentle rolls along the dip are acceptable if the local footwall dip everywhere exceeds 0.78 to 0.87 rad (45 to 50°). Off-dip hanging wall and/or footwall splits can generally be mined selectively from a conventional shrink stope as they are encountered without ad¬versely affecting subsequent continuation of shrinkage mining updip on the main vein. Vertical offsets or major rolls along the dip which cannot be "smoothed over" generally require that a sublevel be established with new draw control development. Blasthole shrinkage methods are much less flexible (and thus less selective) in their ability to accommodate any of these irregularities. Ground Conditions The wall rock must be strong enough to stand with the minimal support afforded by the dynamic mass of broken ore in the stope. During active mining, local sloughing from the walls is restrained, but the broken ore affords little, if any, useful resistance to closure of the stope walls. Such squeezing, if present, may bind up the stope and cause the loss of much ore. Pillars left between and/or within stopes are effective in preventing closure but reduce overall recovery. Walls may be re¬inforced by bolting after each stope cut in conventional shrinkage but not in blasthole shrinkage. Ore in place must be strong enough to stand with no natural support across the stope width, although tem¬porary artificial support or reinforcement may be used locally in conventional stopes. Some spalling or sloughing is permissible in blasthole shrinkage as men are never present in the stope. Physical and/or mineralogical characteristics of the broken ore may impose restrictions on stope design and/or operational plan¬ning, and may even preclude the use of shrinkage al¬together. Examples include: ores which, when broken, are cohesive or which tend to pack or cement together under the influence of ground water, wall pressure, and/ or chemical reaction. Such conditions precipitate er¬ratic draw during mining and often result in difficult and/or incomplete final draw; pyritic ores which oxidize very rapidly in the stopes and may generate heat, imposing a fire hazard by spontaneous combustion; sulfide ores which oxidize sufficiently in the stopes to adversely affect mill recovery by flotation; and ores (es¬pecially those containing uranium minerals) which ex¬ude radon gas and thereby impose ventilation constraints on stope design. In most cases these problems can be minimized by limiting the size of stopes, by minimizing the duration of mining activity in each stope, and by promptly drawing each stope empty following comple¬tion of mining.

Jan 1, 1982

By L. G. Sobral, J. E. Hoffman, T. J. Keefe, A. J. Monhemius, O&apos

The kinetics of electrodeposition of gold from dilute solutions of alkaline KAu(CN)2 at different gold and free cyanide concentrations on a gold rotating disk electrode were investigated by use of a steady-state linear potential scan technique. The design and performance of a laboratory electrowinning cell with a reticulated nickel cathode, operating in the flow-through mode, for the removal of low levels of gold ions from aqueous cyanide media is described. During the batch recirculation of oxygen-free gold cyanide solutions, the gold concentration was reduced from 10 ppm to <0.5 ppm with current efficiencies of up to 80%. The influence of solution flow rate through the cathode and of the pore size of the reticulated nickel foam is discussed. The removal of gold ions from solutions of low ionic strength was also investigated.

Jan 1, 1993

By T. M. Barton

This paper presents several case studies in which a mechanics-based boundary-element program is used to back-calculate the surface subsidence associated with various panels at several northern Appalachian coal mines. The program used in this case study is called LAMODEL, which incorporates a frictionless, laminated overburden into a general-purpose displacement-discontinuity code primarily designed for calculating the stresses and displacements in coal mines or other thin-seam or vein-type deposits. In this paper, the program is used to calculate both the underground convergence and the resulting surface subsidence at five longwall panels and a room-and-pillar section. The fitted subsidence from the model is compared with the field measurements and analyzed. The results from this work show that the LAMODEL program is not as accurate as available empirical subsidence-predictive methods; the expected correlation between the geology and the optimum input parameters is not evident. However, for a mechanics-based program, LAMODEL does provide moderately accurate subsidence calculations, and it is one of a few programs that can even attempt to practically calculate both underground stress and convergence and the resulting surface subsidence.

Jan 1, 2000

By Garland R. Dever

Extensive deposits of high-calcium limestone occur in the Upper Mississippian Warsaw and Ste. Genevieve Limestones in the region of the lower Cumberland, Tennessee, and Ohio Rivers in western Kentucky. Analyses of foot-by-foot samples from 20 quarries and cores in 8 counties show variations between the chemical composition of high-calcium deposits in the 2 formations. Knowledge of probable variations in composition should aid in prospecting for high-calcium stone for a specific use and in evaluating potential deposits. High-calcium strata in the Warsaw consist of bryozoan- and crinoid-rich bioclastic limestones, and in the Ste. Genevieve, predominantly oolitic limestones. Values used for comparison were obtained by averaging the foot-by-foot analyses of each high-calcium deposit, 10 in the Warsaw and 15 in the Ste. Genevieve. Average MgCO3 and Fe2O3 contents are generally lower in the Warsaw than in the Ste. Genevieve. Average SO2, S, and P contents are- generally higher in the Warsaw. No characteristic differences are apparent between the average CaCO3 and A1203 contents of the two formations. The higher S and P contents of the Warsaw are believed attributable to the predominance of biogenic material in these limestones. Within the Warsaw Limestone, higher average S and A1203 contents and lower average Fe2O3 contents were found in deposits in the northern half of the region. In the Ste. Genevieve, vertical variations in composition occur within individual high-calcium deposits containing oolitic and nonoolitic limestones. Lower average CaCO3 contents and higher average contents for the other constituents were found in the nonoolitic intervals.

Jan 1, 1972

By L. A. Taylor, R. O. Colson, D. S. McKay

The production of oxygen on the Moon utilizing indigenous materials will be essential for successful lunar colonization. Several processes have been put forth to accomplish Lunar Liquid Oxygen (LLOX) generation. The schemes that have received the most study to date involve: 1) Reduction of ilmenite (FeTi03) by H2, CO, CH4, CO-Cl2 and Cl2, plasma. 2) Molten silicate electrolysis, both unadulterated and fluoride fluxed. 3) Several others, including H, reduction of glass, carbo-chlorination, HF acid leaching, fluorine extraction, magma oxidation, and vapor pyrolysis. The H2 reduction of ilmenite and molten silicate electrolysis processes have received the most study to date. They are both feasible schemes and have various pros and cons, including their feedstock requirements. However, all processes should be addressed at the onset of LLOX research. This paper reviews several of the most studied and feasible processes for winning oxygen front lunar materials. There is an obvious need for more experimentation and study.

Jan 1, 1994

By Paul C. Schmall

In recent years, several geotechnical methods have been successfully applied to ash pond closure. These include modified dewatering techniques with wellpoints and wells to physically drain the material, soil mixing to create a soil-concrete material for improved accessways as well as improved ground structures within ponds to act as barriers or gravity dams, and geotechnical instrumentation to serve as an early warning of the potential for liquefaction. Given the parallels between fly ash and sluiced in place mine tailings, the application of fly ash remediation method to mine tailings shows significant potential, as illustrated in this paper.

Feb 1, 2024

By M. F. Dibble

Borehole mining, as a viable concept for recovering deep sedimentary mineral resources, historically has been long on promise but short on results. Application of the technology seems best suited for recovery of Miocene phosphate resources on the East Coast of the US. These deposits are potentially enormous in scope and underlie sections of the coastal plain and the near offshore continental shelf. The landside portion of these deposits is generally too deep to mine using conventional open pit methods. Recovery of these minerals requires development of alternative mining systems. A significant proportion of the offshore resources appear to occur at depths below the practical limits of cutter suction dredges and may even exceed the limits of bucket ladder dredges. In most cases, these deposits are covered by overburden that must he removed and redeposited before the mineral section can be mined. The depth, thickness, composition of overburden and environmental sensitivity of a particular location would greatly impact mining system selection. It is probable that dredging would not be an environmentally acceptable mining method in many of the areas underlain by this phosphate rock. In cases such as these, new processes such as borehole mining must be evolved or the resource will remain undeveloped. A borehole mining test program, using an existing prototype borehole system, successfully extracted Miocene phosphate in northeast Florida in 1984-1985. The top of the main mineral zone was 80 m (261 ft) below the surface. Based on the results of this test program. plus subsequent laboratory testing and engineering studies, a new prototype borehole mining system was designed in 1986. This design incorporated several innovations of offering potentially higher productivity and lower power consumption than was achieved with the previous test unit. Economic restraints at that tine prevented the actual construction and testing of this new design.

Jan 1, 1992

By John Frostiak, Don Welch, David Landriault

The use of paste technology for underground cemented backfill is widely accepted in the mining industry today as a potentially cost effective alternative to cemented rock and hydraulic slurry fill. While paste backfill is not the appropriate backfilling method for all mines, there are presently 23 operating tailings paste backfill plants throughout the world. This new technology has gained increasing acceptance as a backfill method over the last decade and its use is now expanding into the surface disposal of tailings using the stacked, thickened tailings disposal concept. Over the next 18 months there will be ten new paste plants commissioned in the world. Five of these ten plants will be used exclusively for surface tailings disposal or a combination of surface disposal and underground backfill production. The thickened tailings disposal concept, developed by Dr. Eli Robinsky in the early 1970&apos;s, has been in use for decades at several mine sites throughout the world (Kidd Creek Mine site in Northern Ontario), but always utilizing non segregating high density tailings slurry central discharge, stacked tailings placement. With the use of a higher solids content paste, the typical 2-degree beach angle can be increased to ten degrees or more with no runoff water and minimal shrinkage crack development through desiccation. This paper presents the design of the first mine to implement true surface paste disposal, Barrick Gold Corp&apos;s Bulyanhulu Mine in Tanzania. The paste system will be commissioned in the first quarter of 2001 allowing performance data to be collected to define the full potential of paste technology for surface tailings disposal.

Jan 1, 2001

By K. Biswas, P. Tsang, S. S. Peng

A survey was undertaken in late 1993 and early 1994 to gain insight into the current status of pillar-design practice in underground coal mines in the United States. A total of 552 room-and-pillar mines and 75 longwall mines were surveyed to collect general, geological and mine-plan information. The collected information was critically analyzed with special emphasis on the pillar-design criteria adopted in these mines. A safety factor analysis was also carried out for the pillars in the room-and-pillar mines.

Jan 1, 1997

Mining has traditionally been seen as a relatively risky business. Exploration is expensive and the failure rate is high, while start-up costs commonly run to many hundreds of million dollars. To justify the risk and expense, the returns from successful projects must also be high. Accurate estimation of mineral resources and reserves is fundamental to the processes of optimising returns and managing risk in mining developments. Systematic development from initial exploration through a series of development stages enables expenditure and risk to be balanced. As a consequence, large scale modern mine are more akin to manufacturing businesses. This paper examines the life history of a mineral reserve estimate from the perspective of a large mining company against a background of modern estimation and reporting requirements. It will also touch on what constitutes, in the author&apos;s opinion, `best&apos; practice in reserve estimation, and how this concept can be spread more widely. [Abstract to be re-written and reduced to 100 words]

Jan 1, 2003

By J. H. Tylczak, R. Blickensderfer, D. J. Singleton

As part of its research on wear of mining and mineral processing equipment, the US Bureau of Mines evaluated the erosion-corrosion characteristics of 12 alloys in phosphoric acid waste water. Tests were conducted on 2- and 5-cm-diam specimen8 in 12- and 60-cm-diam ball mills by grinding phosphate rock in gypsum pond water with an initial pH of 1.6. For comparison, grinding tests also were conducted in tap water. Static corrosion tests in gypsum pond water also were carried out. For all alloys tested, the wear by erosion-corrosion was greater than the sum of erosion plus static corrosion. A low-alloy, high-carbon steel and a high-manganese, nitrided stainless steel were the most cost-effective materials

Jan 1, 1987

By R. Harris, M. Misra

Column flotation technology has a demonstrated potential for recovering clean coal from high-ash waste fines. The effect of different variables such as air flow rate, frother addition, depressant addition, percent solids and particle size on the quality of separation are discussed. Results indicate that air flow rate and depressant addition are the important variables which control the selectivity in the flotation of fine coal in a column cell. Separation efficiency achieved with column cells is significantly better than that with conventional flotation cells. Under appropriate experimental conditions with a deep column, more than 80 percent coal can be recovered in the concentrate containing 10 percent ash from waste fines containing over 60 percent ash.

Jan 1, 1988

By L. J. Prosser, A. T. Iannacchione

From 1991 through 1995, 44 miners out of a total workforce of less than 2,000 were fatally injured in the stone industry. Of these, 12 occurred at underground mining operations with nine deaths resulting from roof or rib falls. A safer environment can be achieved by evaluating the nature of the hazardous ground and by developing more efficient and effective ground-control strategies. Roof and rib conditions were observed and assessed in 33 underground stone mines in Illinois, Indiana, Kentucky, Maryland, Missouri, Pennsylvania and West Virginia (Fig. 1). Hazard assessment indicated that the ground failures that occurred under moderate to substantial overburden, i.e., >30 m (100 ft), were caused by stress concentrations and geologic structures. Ground failures near the surface are caused by solution (water) processes. Selection of the mining horizon and mine-layout decisions

Jan 1, 1999

By S. S. Peng, V. Z. Sun

One of the major concerns of ground control in a coal mine is the prevention of roof, pillar and floor failures of the mine entry. To better design and evaluate the stability of a mine entry within a certain level of confidence, a reliability approach is introduced. The advantage of using the reliability approach, rather than the conventional safety-factor (SF) method, is that it not only takes variability of both rock/coal properties and stress into consideration but also assesses the stability of each component and the mine entry as a whole. Therefore, it is anticipated that by using this approach the current ground-control technology in a coal mine will convey a better sense of confidence. The results of a case study show that there is a significant difference in mine entry stability between the current design method and the reliability approach.

Jan 1, 1999