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By Ken Deter

I have some general comments on optimizing profitability based on working in heap leach operations in tropical environments. Many of my planned comments this morning have been stated earlier in the Conference. I get the opportunity to say only a few words, and follow up on some specific comments and concerns. In normal operations, stack the ore higher, where possible. The surface area of the stacked ore controls the total solution volume that can be sprayed on the pad at a specific time period. With higher lifts, solution will contact more ore particles as it percolate through the lift, and therefore the solution grade flowing through the base will be increased, whether it is freshly stacked or almost completely leached ore. This also allows more tonnage to be stacked per lift on the pad, decreasing conveyor movement and downtime. Increasing from 6 to 8 meter (20 to 26 feet) lifts have worked successfully, and some mines are using 10 meter (33 feet) lifts in 10,000-15,000 dmtpd plants. The same concept will work with smaller plants, but initial 6 meter lifts should be used until the start-up period has shown the leaching is working as planned. As the stacked ore lifts get higher, the solution flowing through the highest lift is diluted by solution retained in the lower ones. Eventually no real increase is noted from putting fresh ore in service, and gold production rate suffers. To address the dilution problem, an HDPE liner should be installed between each 2 to 4 lifts, depending on each site specific lift height, recovery, % slump in the lift, recovery rates, and grade treated, among others. The surface of the previous lift should be leveled and sloped to control the effluent draining from the planned lift, then the liner is placed, including berms and discharge pipes. The intermediate HDPE liner does not need to be completely sealed as would be the case with the base liner, and can accept some solution penetrating into the lower lifts. A one-mm (40-mil) thickness liner can be used. This intermediate liner allows the upper lift(s) higher-grade effluent to report to the pad discharge point. This will maintain higher gold solution grades going to the ponds. The slopes of the previous lifts should not be covered with HDPE in this concept, and precipitation falling on the lower lift slopes and benches can be directed to barren or surge ponds.

Jan 1, 1999

By Victor C. Bryan

One of the most sought after answers in any consideration of blasting activities relative to mines and quarries is how to safely and efficiently use explosives. While that discussion might be relatively straightfonvard in a “site specific” situation. it becomes more complex when viewed from a broad geographic perspective. Such is the case in considering the operations of one United States producer of crushed stone with surface and underground mines spread over a thirteen state area. Differing geologic environments set conditions that dictate to the blaster as now specifically to get optimum results from the blast hole. This treatise is a discussion of the various methods, materials, and principles used in conducting such blasting operations.

Jan 1, 1991

By Justin Reed, Sara Bewley, Dwight M. Kinnes, Maryam Sadeghi, Ed Hart

This study compares wet gravity and dry magnetic separation for preconcentrating an allanite-rich rare earth ore from the Halleck Creek deposit. Material in the –300/+86 μm size range was treated to define a flowsheet that maximizes total rare earth oxide (TREO) recovery. The ore is dominated by feldspar and quartz, with hastingsite, an iron-bearing aluminosilicate. Because hastingsite’s density and magnetic response closely match allanite, it is difficult to reject using conventional methods. Two separation approaches were evaluated: MG12 spirals for gravity concentration and an Induced Roll Magnetic Separator (IRMS) for dry magnetic upgrading. MG12 spirals performed well in the rougher stage but provided limited gains when used as cleaners. When IRMS was applied as the cleaner stage, TREO grade and recovery improved substantially, showing that a dry magnetic cleaner circuit can effectively replace the wet gravity cleaner step. MG12 spirals remain suitable for highcapacity rougher duty, but dry magnetic separation delivers higher selectivity, simpler operation, and eliminates water use—offering a more efficient and sustainable preconcentration route in water-constrained environments.

Feb 22, 2026

By P. Todd, K. Dever, P. S. Mills, A. A. Campoli

Resin bolt performance is dependent on the size of the resin annulus, the difference between the hole and bolt radii. The vast majority of the millions of resin roof bolts installed each year in the United States employ an annulus of 0.125 or 0.188 inches. Short encapsulation pull tests with resin annuluses of 0.125 and 0.095 inch resisted an average of 2.25tons per grouted inch, with no significant variation. Resin loses to simulated fractures increased as annulus size decreased, with a minimum loss of35% and a maximum loss of 60% associated with0.188 and 0.095 inch annuluses, respectively.

Jan 1, 2000

Optimal gold recoveries largely depend on efficient liberation of the gold or gold-bearing minerals as well as on maximum access for the lixiviant. Thus, crushing and grinding may become the most costly part of gold ore processing. It may, in fact, represent 50% to 70% of a plant's operating costs (Michaelis, 1989). The increasing significance of refractory gold will require even more cost-efficient comminution of ultrafine sulfides and/or native gold. Transitional ore zones with semi refractory gold may respond to direct leaching if better particle fracturing is being achieved (Baum, 1991, 1993a). Many oxide ores have been left in the ground because they were unsuitable for heap leaching since conventional comminution could not facilitate economic gold extraction. Also, the costs of conventional bal1 mill operations did not permit the recovery of gold from many ultrafine, low-grade refractory ores. As shown by Marsden et al. (1993), the trend to treat ores with progressively lower grades will involve power-and leach-efficient grinding applications. Our current laboratory work shows that high-pressure grinding roll (HPGR) Polycom grinding may enable many mining companies to add large amounts of these "non-leachable" ores to their current heap or mill leach feeds. This finding may dramatically change the economics of many oxide gold ore operations and may add significant reserves to ores amenable to direct cyanidation.

Jan 1, 1994

Many of the deeper ore reserves in current gold operations as well as new gold deposits consist of refractory ore types. As process development for these mineralizations remains time-consuming and costly, minimization of exploratory testwork and the early identification of feasible process routes is critical. Since most of the processing parameters are related to the mineralogy of the ore (Marsden & House, 1992), thorough mineralogical analyses during exploration and development phases have become a technical and economical requirement. The determination of mineralogy-related processing parameters and especially the gold mineralogy is even more important for these ore types because refractory gold ores consist of many transitions from partially leachable to totally refractory. In addition to sulfide and gangue encapsulation, various other refractory conditions may exist. Together with accessory mineralogical detriments (e.g., alteration/ slimes, reagent consumers, scale formation, mineral preg robbing) they may complicate process selection and hamper optimal plant operation.

Jan 1, 1993

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

No method of roof control yet devised has proven to be universally acceptable for the wide range of strata conditions experienced in U. S. coal mines. However, a relatively new innovation, the roof truss, has proven itself capable of preventing roof failures which both conventional and resin bolting could not. Presently, the economics of roof trusses are not favorable when compared to bolting due primarily to drilling and installation procedures, but this is expected to change with improved truss and equipment designs. To realize the full potential of this new method of roof control, design criteria are needed which can optimize the effect of the truss for varying strata conditions.

Jan 6, 1978

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

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

Jan 1, 1979

By Neall. George M.

No method of roof control yet devised has proven to be universally acceptable for the wide range of strata conditions experienced in U.S. coal mines. However, a relatively new innovation, the roof truss, has proven itself capable of preventing roof failures which both conventional and resin bolting could not. Presently, the economics of roof trusses are not favorable when compared to bolting due primarily to drilling and installation procedures, but this is expected to change with improved truss and equipment designs. To realize the full potential of this new method of roof control, design criteria are needed which can optimize the effect of the truss for varying strata conditions. A review of the literature reveals that some work has been conducted on the roof truss. 1 Gambrell and Haynes (1970) showed that the roof truss is superior to angled roofbolts installed through a header. Sheorey, Verma, and singh2 (1973) used statics to prove that the optimum angle of inclination should be greater than 450? However, to date, no systematic attempt has been made to establish design criteria. In this report, the problem was approached in three ways. First, model studies were performed to determine general trends of truss support as a function of a number of variables. This provided a better understanding of the support mechanism involved. Secondly, beam theory was employed to calculate the actual support capacity of a truss, realizing the many simplifications, assumptions, and shortcomings of this approach to such a complicated problem. Last, field testing will be utilized to test the hypotheses previously developed and provide hard data to further synthesize roof truss design criteria. Field testing has recently commenced and should be complete before mid 1977.

Jan 1, 1977

By Lawrence Adler

The authors are to be complimented on using a novel and potentially helpful technique to analyze the effects of mine roof trusses. However, some criticisms of their application and analysis seem appropriate to further its acceptance. This discussion will deal with the normal structural analysis format covering: 1) material, 2) load, 3) structure, and 4) analysis. It will then briefly review the testing procedures and finally cover miscellaneous items.

Jan 12, 1978

By George P. Jr Schivley

Data is presented showing Penetration Rate (PR) versus Force-on-the Bit (FB) and Bit Angular Speed (N). Using this data, it is shown how FB and N each uniquely contribute to the PR for any particular drilling situation. This data represents many mining situations; including coal, copper, gold, iron ore and limestone quarrying. The important relationship between Penetration per Revolution (P/R) and the height of the cutting elements of the bit (CH) is discussed. Drill performance is then reviewed, considering the effect of FB and N on bit life. All this leads to recommendations for the operating values of FB and N for drilling situations where the rock is not highly abrasive and bit replacements are because of catastrophic failure of the bit cone bearings.

Jan 1, 1995

By Federico Chavez

The process for recovering lead from battery residues used in several Mexican Plants include lead blast furnace and rotary furnaces. The formers are basically fed with recycled batteries together with the necessary additives to produce metallic lead. One of the by-products generated in the blast furnace is a dust mainly formed by lead sulfide. The rotary furnace is used to recover lead .from lead batteries, dross from the refining-alloying process and the leady fines produced in the blast furnace. This paper describes thennodynamic analysis made to the rotary furnace process to understand the effect that each component of the system has upon the lead recovery. In order to enhance the process, the influence of iron chips composition and excess coke on slag viscosity is discussed. A computer optimization program . has been developed to design the furnace charge based on material and energy balances, .together with the thennodynamic equilibrium calculation.

Jan 1, 1995

By Ivan C. Machado

In full autogenous grinding (FAG Mills) the ore itself determines almost exclusively the product size that will be produced, as opposed to conventional steel-media grinding where the product size can be "engineered" by changing media size, feed size, feed rate, classifier cutpoint size, etc. SAG Mills are in the middle of the way, since they offer some room to engineer their product size by properly defining some critical process variables, such as feed size, ball charge and size, mill charge and speed. SAG Mill product P80 values are typically in the range of 500 µm (35 US mesh) to 2000 µm (10 US mesh.) Whenever it is possible to determine an optimum value for the SAG Mill product P80 value (which will be the Ball Mill fresh feed F80 value), the Engineer should define the process variables that will yield a SAG product size as close as possible to the optimum calculated P80 value. In general, a coarser SAG product will draw less power from the SAG Mill and more power from the Ball Mill, and conversely, a finer SAG product will draw more power from the SAG Mill and less power from the Ball Mill. Metal consumption analyses also have to be taken into consideration, for they typically account for 30% to 45% of the total grinding cost. This presentation addresses the issue of defining an optimum SAG Mill product size (X=P80), during the design phase of the SAG and Ball milling circuits, from a general perspective, i.e., the SAG product size that will lead to an overall SAG Mill - Ball Mill optimum performance. Bond's Third Law approach is used, as well as the SPI (SAG Power Index), and also the generally accepted equations and Engineering Practice currently utilized to design SAG Mill

Jan 1, 1999

By Thomas M. Barczak

The 1990s brought an unprecedented increase in the development of innovative technologies to provide more effective and easier-to-install roof support in underground mines. To facilitate the application of these technologies in improving mine safety, the National Institute for Occupational Safety and Health (NIOSH) developed the Support Technology Optimization Program (STOP). STOP is a Windows-based software program that provides mine operators with a simple and practical tool to make engineering decisions about the selection and placement of these various roof support technologies. The program includes a complete database of the support characteristics and loading profiles obtained through safety performance testing of these supports at the NIOSH Safety Structures Testing Laboratory. A support design criterion in the form of the required support load density at a specified convergence can be established from four options: (1) a database of measured ground reaction obtained from various mines or ground behavior information input by the user,(2) load requirements based on a detached roof block or rock failure height, (3) criteria based on the current roof support system, and/or (4) arbitrary criteria set by the user. Using these design criteria, the program will determine the installation requirements for a particular support technology that will provide the necessary support load density and convergence control. Optimization routines are also available to determine the most efficient support design for a user-specified support installation. In addition to these performance measures, STOP can be used to compare material handling requirements and installation costs. Comparisons among the various support technologies are easily made, including a graphical analysis of relevant support parameters. This paper describes STOP and its application to optimizing standing secondary roof support systems.

Jan 10, 2000

By Thomas M. Barczak

The decade of the 90's brought an unprecedented increase in the development of innovative technologies to provide more effective and easier to install roof support in underground mines. To facilitate the application of these technologies to improve mine safety. National Institute for Occupational Safety and Health (NIOSH) developed the Support Technology Optimization Program (STOP). STOP is a windows based software program that provides mine operators with a simple and practical tool to make engineering decisions regarding the selection and placement strategy of these various standing roof support technologies. This program includes a complete data base of the support characteristics and loading profiles obtained through safety performance testing of these supports at the NIOSH Safety Structures Testing Laboratory. Support design criteria in the form of the required support load density at a specified convergence can he established from four options: ( 1 ) a data base of measured ground reaction data from various mines or ground behavior information inputted by the user. (2) determination of loading requirements based on a detached roof block or rock failure height. or (3) criteria based on the current roof support system, and (4) arbitrary criteria set the by user. Using these design criteria, the program will determine the installation requirements for a particular support technology that will provide the necessary support load density and convergence control. Optimization routines are also available to determine the most efficient support design for a user specified support installation- In addition to these performance measures. STOP can be used to compare material handling requirements and installation costs. Comparisons among the various support technologies are easily made including graphical analysis of relevant support parameters. This paper describes the STOP and its application to optimize standing secondary roof support systems.

Jan 1, 2000

A model was developed to facilitate secondary tailgate support selection based on analysis of over 100 case studies, compiled from two different surveys of operating longwall coal mines in the United States. The ALPS (Analysis of Longwall Pillar Stability) program was used to determine adequacy of pillar design for the successful longwall case histories. A relationship was developed between the secondary support density necessary to maintain a stable tailgate entry during mining and the CMRR (Coal Mine Roof Rating). This relationship defines the lower bound of secondary support density currently used in longwall mines. The model used only successful tailgate case history data, with adequate ALPS SF according to the CMRR for each case. This model facilitates mine design by predicting secondary support density required for a tailgate entry depending on the ALPS SF and CMRR, which can result in significant economic benefits.

Jan 1, 1996

By W. J. Kempson

As a result of the rising electrical energy costs in South Africa, a method was sought to reduce the overall electrical consumption of typical shaft systems. A typical shaft configuration was analysed and the primary energy consumers were identified. The ventilation fans for this system were found to consume 15% of the total energy of the shaft system. It was calculated that more than 50% of this energy is consumed by the shaft itself, more specifically, by the pressure losses that occur in the shaft as the ventilation air passes through it. In order to ensure that the theory being used for the evaluation of these shaft systems is accurate, a total of five shafts were instrumented and the actual pressure losses over the shafts plotted against time. These shafts were then analysed from a theoretical perspective. Finally, in order to ensure a thorough understanding of the behaviour of the ventilation air in shaft systems, the systems were simulated using computational fluid dynamics (CFD) techniques. There were significant discrepancies between the theoretical analysis and the CFD simulation during the initial comparisons. This discrepancy lessened as the complexity of the CFD models increased, until when the complete shaft was modelled using the full bunton sets, the pipes, and the flanges, the difference between the theoretical evaluation and the CFD simulation was small. This result demonstrates that the theory is insufficient and that the interrelated effect of the buntons and fittings has not been fully appreciated by current theory. The final phase of the work presented here was to evaluate the costeffectiveness of using different bunton shapes and shaft configurations. It is shown that the increase in the pressure losses and therefore the direct operating costs of the shaft can vary by as much as 80%, depending on the bunton configuration chosen. The placement of the piping in the shaft can increase the pressure losses, and therefore the direct operating costs of the shaft, by as much as 12%, depending on the placement of the piping in the shaft; this effect includes the use of flanges. The use of fairings on a large cage can reduce the resistance that the cage offers to the ventilation flow by as much as 30%. This, however, does not translate into a direct saving because as the cage moves through the shaft, the overall effect is transitory. These savings can be significant when the items highlighted in this work are applied correctly. Keywords shaft design, ventilation, energy consumption, modelling, CFD.

Dec 1, 2013

By Bien Ferrer

In most smelting operations, there is a strong incentive to improve safety, reduce downtime and extend the campaigns of furnaces and related equipment. An asset management strategy that is based on information, rather than simple throughput metrics and intuition, can be developed to improve uptime and unlock value currently unrealized at existing smelters. Advances in digital technology, including data analysis, modelling and monitoring for hot pyro-metallurgical vessels, make it possible to assess the current condition of an asset, as well as to make more informed projections about its future condition, using on-line and historical data. This would allow practices around asset management to be more proactive and less reactive, with the end-result of optimized uptime. ‘Digital Asset Management’ describes a concept in which digital technology is used to support an asset management strategy and accelerate decision-making by providing information on what, why and how an asset should be designed, operated or maintained. With the overall goals of effective resource utilization, optimized operations, capital efficiency and social acceptance, Smelter 4.0 is a program of holistic improvement and is underpinned by the effective use of data to drive decisions. This paper focuses on Asset Reliability, which is one of the improvement pillars of Smelter 4.0, and introduces a methodology that is aimed at closing the gap between recorded data and evidence-based decision-making through Digital Asset Management.

By Ali Hashemi, Anna P. Perry

The PQ Deeps (PQD) is a production zone located at Newmont’s Musselwhite Mine in northern Ontario. Development started for the PQD in late 2017, and the first stopes were mined starting in early 2021. Unlike most historic longitudinal stoping of narrower orebodies at Musselwhite, the PQD is a wider orebody following the plunge of a fold hinge and is extracted primarily by transverse stoping. Some geotechnical challenges were experienced during early stoping efforts, necessitating design changes throughout late 2021 and early 2022 to optimize the stope design. This work describes the mining methods and design changes to transverse PQD stoping based on geotechnical lessons learned during early stoping efforts.

Jan 1, 2023

By Shahé Shnorhokian

Numerical modeling is increasingly being used as a tool to examine complex geomechanical issues related to rockmass behaviour and stability of underground openings. The in-situ stress regime and the evolution of mining-induced stresses during the life of an operation, taking into account the planned stope sequence, are topics that can only be handled through the use of a realistic and calibrated numerical model. This paper presents a case study of a deep underground Canadian mine in northern Ontario. A FLAC3D model was constructed and calibrated based on in-situ stress measurements. In the first stage, the evolution of stress regime was studied and compared with the occurrence of rockburst events during a period of three years. The locations of clusters were examined and compared to the mining sequence during that period. In the second stage, stresses within a diminishing ore pillar were examined based on the planned stope sequence and compared to levels deemed acceptable to the mine. Based on the results, an alternative sequence was arrived at to reduce stress accumulation in the pillar, or induce it to occur at a further distance from a brittle dyke in the vicinity of the orebody.

Aug 1, 2013