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This dissertation will explore the process followed in the design of a sub-level open stope (SLOS) by using examples of actual stopes scheduled to be in production between August 2006 and February 2007. The main objective is to give the reader an understanding into sub-level open stoping and the design process followed. The objective here is to present a design methodology applicable to sublevel open stoping, but also to then bridge the gap between theory and practice by applying said methodology to an actual design example. The design examples used in this dissertation is based on the O640, L651 and N659 stopes in the 3000 Orebody of Xstrata Copper Operation’s Mount Isa Mine, located in North-West Queensland, Australia. The actual design reports as required by the mine are attached in Annexure 1 through 3. Given the similarities of the designs, only O640 will be analysed comprehensively within the main content of this report, with L651 and N659 discussed specifically insofar issues that were unique to these stopes. With the design of O640, all aspects or design considerations as stipulated in the design process were discussed and analysed so as to define the final stope shape. These design considerations include: <ul> <li> Faulting</li> <li> Grade Contours</li> <li> Existing Development</li> <li> Surrounding Fill masses</li> <li> Rock Mechanics</li> </ul> Once the the final stope shape has been set, options regarding stope extraction will take place. This is where the initial stope layout takes place and where the engineer looks at the advantages and disadvantages of all the different options available in mining the stope. In this phase, the most effective extraction option is decided upon. Once the engineer have decided a final stope shape and extraction option, the stope will be analysed in further detail referring to drilling, the amount of drawpoints, ventilation and other stoping requirements. These are all defined as stope design features and are considered a general summary of the stope design. The design features phase is closely followed with all the safety considerations that have been taken into account since the stope design started. Main concerns and stope specific safety issues are discussed and possible solutions given. It is part of the work of the mine planning engineer to anticipate all possible safety issues and make the production department aware of what can be expected during the development, mining and filling activities of every stope. At this stage the design of the stope nears completion. The remainder of the design now goes into more detail and addresses the critical tasks that from part of sub-level open stoping. These include: <ul> <li> Reserves and Scheduling</li> <li> Development and Drilling</li> <li> Production and Firing (Blasting)</li> <li> Ventilation</li> <li> Services</li> <li> Filling</li> <li> Economic Analysis</li> </ul> Although all of the abovementioned have already been mentioned during the design features phase, it is still required to give additional details so the different departments involved have an accurate idea of what to expect, when to expect it and therefore be able to sufficiently plan for it. It must be noted that it does happen that something may be “discovered” during any stage of the final design, which may render the current design undesirable. When this happens the stope must be re-designed until all issues have been resolved or at the least have been managed appropriately. Even though this report does not go into detail with the L651 and N659 designs, these designs are included as they bring to light issues that may arise that are unique to individual stopes. L651 looks at how a design drastically changes when ore not planned for is discovered. N659 looks at what happens when a stope is the first to be mined in an area with inadequate infrastructure. The main content of the dissertation discusses and explains the design procedure as it would take place at Mount Isa Mines, but it is still quite difficult to follow logically. For this reason a flowchart was included to give the reader a more comprehensive summary of the design process. / Dissertation (MEng)--University of Pretoria, 2011. / Mining Engineering / unrestricted

No description available.

Over the past three decades, technological innovations with respect to cemented paste backfill (CPB) as a means of ground support has allowed for increased production within the mining industry, management mine waste costs, as well as the improvement of the overall health and safety of underground mining operations. Despite the extensive use of this relatively new ground support material, many fundamental factors affecting the design of safe and economical CPB structures are still not well understood.Recently, a significant amount of academic and industry research has been conducted to better understanding the distribution of stress with respect to primary-secondary extraction sequencing for stope-and-fill mining operations. While current, as well as past research, as provided a wealth of knowledge on the distribution of stress through the fill material itself, it lacks in providing an examination into the mechanism by which stress is able to redistribute itself through the backfill material as well as within the surrounding rockmass. The scope of this work is to optimize stope-and-fill extraction sequencing through the analysis of stress distributions as well as local and global stability of multiple narrow verticalfully-drained backfilled stopes. Scientific investigations into the behavior of the CPB material and surrounding rockmass will result in animproved understanding of how to better implement engineered paste-fill materials as a means of ground support for underground mining operations. Numerical simulations (FLAC3D and RocScience) were utilized in analyzing hypothetical (literature) as well as site-specific (field) case studies. While these simulations confirm generalized stress behaviors within the backfill material for single and adjacent stopes, stress redistributions within the surrounding rockmass as well as the rock-pillarindicate the development of tensile and compressive zones. From these results, one is able to better approximate ground and CPB instability with respect to site-specific conditions, geometries, and material properties. These simulations have been validated with respect to published analytical solutions, numerical simulations, and site-measurements for single (isolated) and adjacent narrow vertical fully-drained backfilled stopes.

This thesis presents new methods to improve open stope hangingwall (HW) design based on knowledge gained from site visits, observations, and data collection at underground mines in Canada, Australia, and Kazakhstan. The data for analysis was collected during 2 months of research at the Hudson Bay Mining and Smelting Ltd. Callinan Mine in Flin Flon, Manitoba, a few trips to the Cameco Rabbit Lake mine in northern Saskatchewan, and 3 years of research and employment at the Xstrata Zinc George Fisher mine near Mount Isa, Queensland, Australia. Other sites visited, where substantial stope stability knowledge was accessed include the Inco Thompson mines in northern Manitoba; BHP Cannington mine, Xstrata Zinc Lead Mine, and Xstrata Copper Enterprise Mine, in Queensland, Australia; and the Kazzinc Maleevskiy Mine in north-eastern Kazakhstan. An improved understanding of stability and design of open stope HWs was developed based on: 1) Three years of data collection from various rock masses and mining geometries to develop new sets of design lines for an existing HW stability assessment method; 2) The consideration of various scales of domains to examine HW rock mass behaviour and development of a new HW stability assessment method; 3) The investigation of the HW failure mechanism using analytical and numerical methods; 4) An examination of the effects of stress, undercutting, faulting, and time on stope HW stability through the presentation of observations and case histories; and 5) Innovative stope design techniques to manage predicted stope HW instability. An observational approach was used for the formulation of the new stope design methodology. To improve mine performance by reducing and/or controlling the HW rock from diluting the ore with non-economic material, the individual stope design methodology included creating vertical HWs, leaving ore skins or chocks where appropriate, and rock mass management. The work contributed to a reduction in annual dilution from 14.4% (2003) to 6.3% (2005), an increase in zinc grade from 7.4% to 8.7%, and increasing production tonnes from 2.1 to 2.6 Mt (Capes et al., 2006).

This thesis presents the results of open stope stability and dilution research which focused on evaluating and quantifying stress, undercutting, blasting and exposure time and their effect on open stope stability and dilution. Open stope mining is the most common method of underground mining in Canada. Unplanned stope dilution is a major cost factor for many mining operations. Significant advances in empirical stability and dilution design methods have improved our ability to predict probable dilution from open stoping operations. However, some of the factors that influence hanging wall dilution are either ignored or assessed in purely subjective terms in existing designs. This thesis attempts to quantify these factors, from a geomechanics perspective, to assist in predicting and minimizing dilution. A comprehensive database was established for this study based on two summers of field work. Site geomechanics rock mass mapping and classification were conducted and case histories were collected from Cavity Monitoring System (CMS) surveyed stopes from Hudson Bay Mining and Smelting Co. Ltd. (HBMS) operations. The stope hanging wall (HW) zone of stress relaxation was quantified based on extensive 2D and 3D numerical modelling. Stress relaxation was linked to the stope geometry and the degree of adjacent mining activity. The influence of undercutting on stope HW stability and dilution was analysed using the case histories collected from HBMS mines. An undercutting factor (UF) was developed to account for the undercutting influence on stope HW dilution. Numerical simulations were conducted to provide a theoretical basis for the undercutting factor. A relationship was observed between the degree of undercutting, expressed by the UF term and the measured dilution. Many factors can significantly and simultaneously affect a blast performance, which may result in blast damage to stope walls. Major blasting factors which influence stope HW stability were identified. The influence of blasting on stope HW stability and dilution was evaluated based on the established database. The HBMS database, Bieniawskis stand-up time graph, as well as Geco mine case histories were used to evaluate the influence of exposure time on stope stability and dilution. Relating increased mining time to increased dilution allows the mining engineer to equate mining delays to dilution costs. Each of the factors assessed in this study was studied independently to assess its influence on stope dilution, based on the HBMS database. The factors influencing dilution often work together, so a multiple parameter regression model was used to analyze the available parameters in the HBMS database. The findings of this research greatly improve an engineers ability to understand and to predict the influence of mining activities and stoping plans on hanging wall dilution.

This thesis presents the results of open stope stability and dilution research which focused on evaluating and quantifying stress, undercutting, blasting and exposure time and their effect on open stope stability and dilution. Open stope mining is the most common method of underground mining in Canada. Unplanned stope dilution is a major cost factor for many mining operations. Significant advances in empirical stability and dilution design methods have improved our ability to predict probable dilution from open stoping operations. However, some of the factors that influence hanging wall dilution are either ignored or assessed in purely subjective terms in existing designs. This thesis attempts to quantify these factors, from a geomechanics perspective, to assist in predicting and minimizing dilution. A comprehensive database was established for this study based on two summers of field work. Site geomechanics rock mass mapping and classification were conducted and case histories were collected from Cavity Monitoring System (CMS) surveyed stopes from Hudson Bay Mining and Smelting Co. Ltd. (HBMS) operations. The stope hanging wall (HW) zone of stress relaxation was quantified based on extensive 2D and 3D numerical modelling. Stress relaxation was linked to the stope geometry and the degree of adjacent mining activity. The influence of undercutting on stope HW stability and dilution was analysed using the case histories collected from HBMS mines. An undercutting factor (UF) was developed to account for the undercutting influence on stope HW dilution. Numerical simulations were conducted to provide a theoretical basis for the undercutting factor. A relationship was observed between the degree of undercutting, expressed by the UF term and the measured dilution. Many factors can significantly and simultaneously affect a blast performance, which may result in blast damage to stope walls. Major blasting factors which influence stope HW stability were identified. The influence of blasting on stope HW stability and dilution was evaluated based on the established database. The HBMS database, Bieniawskis stand-up time graph, as well as Geco mine case histories were used to evaluate the influence of exposure time on stope stability and dilution. Relating increased mining time to increased dilution allows the mining engineer to equate mining delays to dilution costs. Each of the factors assessed in this study was studied independently to assess its influence on stope dilution, based on the HBMS database. The factors influencing dilution often work together, so a multiple parameter regression model was used to analyze the available parameters in the HBMS database. The findings of this research greatly improve an engineers ability to understand and to predict the influence of mining activities and stoping plans on hanging wall dilution.

This thesis presents new methods to improve open stope hangingwall (HW) design based on knowledge gained from site visits, observations, and data collection at underground mines in Canada, Australia, and Kazakhstan. The data for analysis was collected during 2 months of research at the Hudson Bay Mining and Smelting Ltd. Callinan Mine in Flin Flon, Manitoba, a few trips to the Cameco Rabbit Lake mine in northern Saskatchewan, and 3 years of research and employment at the Xstrata Zinc George Fisher mine near Mount Isa, Queensland, Australia. Other sites visited, where substantial stope stability knowledge was accessed include the Inco Thompson mines in northern Manitoba; BHP Cannington mine, Xstrata Zinc Lead Mine, and Xstrata Copper Enterprise Mine, in Queensland, Australia; and the Kazzinc Maleevskiy Mine in north-eastern Kazakhstan. An improved understanding of stability and design of open stope HWs was developed based on: 1) Three years of data collection from various rock masses and mining geometries to develop new sets of design lines for an existing HW stability assessment method; 2) The consideration of various scales of domains to examine HW rock mass behaviour and development of a new HW stability assessment method; 3) The investigation of the HW failure mechanism using analytical and numerical methods; 4) An examination of the effects of stress, undercutting, faulting, and time on stope HW stability through the presentation of observations and case histories; and 5) Innovative stope design techniques to manage predicted stope HW instability. An observational approach was used for the formulation of the new stope design methodology. To improve mine performance by reducing and/or controlling the HW rock from diluting the ore with non-economic material, the individual stope design methodology included creating vertical HWs, leaving ore skins or chocks where appropriate, and rock mass management. The work contributed to a reduction in annual dilution from 14.4% (2003) to 6.3% (2005), an increase in zinc grade from 7.4% to 8.7%, and increasing production tonnes from 2.1 to 2.6 Mt (Capes et al., 2006).

Bacterial leaching utilises bacteria, ubiquitous to sulphide mining environments to oxidise sulphide ores. The sulphide mineral chalcopyrite is the most common copper mineral in the world, comprising the bulk of the known copper reserves. Chalcopyrite is resistant to bacterial leaching and despite research over the last 20-30 years, has not yet been economically bioleached. Attempts have been made to use silver to catalyse the bacterial leaching of chalcopyrite since the early seventies. The majority of reported testwork had been performed on finely ground ore and concentrates in agitated batch reactors. This project used silver to catalyse the bioleaching of chalcopyrite in shake flasks, small columns and large columns. The catalytic effect was extensively studied and experimental parameters were varied to maximise copper recovery. Silver was also used to catalyse the ferric leaching of chalcopyrite at elevated temperatures. It was noted that the leaching performance of chalcopyrite in shake flasks compared to columns was markedly different. The specific differences between shake flasks and columns were qualified and separately tested to determine which parameter(s) affected the bioleaching of chalcopyrite. It was found that the ore to solution ratio, aeration, addition of carbon dioxide, solution distribution and small variations in the leaching temperature did not significantly effect the bioleaching of chalcopyrite ore in columns. The method of silver addition to columns did significantly affect the overall copper extraction. The ore in shake flasks was subjected to abrasion between ore particles and with the base of the flask. A test was designed to mimic the shake flask conditions, without the abrasion. The low abrasion test performed similarly to a column, operated with optimum silver addition. This indicated that the inherent equipment difference between shake flask and column operation largely accounted for the difference in leaching performance. Chalcopyrite ore was biologically leached in large columns. The ore crush size and other conditions were typical of those used in the field. The biological leach achieved 65% copper extraction in 160 days. This level of copper extraction is significantly higher than any previously reported results (typically /10% copper extraction) and represents a significant advance in the bacterial leaching of chalcopyrite ore. Due to the inherent high temperature within underground stopes, it was decided to investigate the possibility of separating the leaching and the bacterial oxidation stages. The concept of separate bacterial and ferric leaching has been previously suggested, however the application to a stope, and heat exchange between the process streams was a novel approach. Large column ferric leaches at 70 oC illustrated the technical feasibility of this process. Copper extraction was rapid and high (70% in 100 days of leaching), even when a reduced level of silver catalysis was used. After leaching in large columns, samples of ore were taken for analysis by optical mineralogy. The analysis gave valuable insights into the nature of reaction passivation on chalcopyrite ore. In particular, it was discovered that the precipitation of goethite was a major limiting factor in the bioleaching and ferric leaching of chalcopyrite in columns. In addition, reduced sulphide species were detected on the surface of residual chalcopyrite, giving an indication of the sequential nature of the chalcopyrite reaction chemistry. The bacterial population was characterised using DNA techniques developed during the project. Qualitative speciation was carried out and compared between the columns, down the columns and over time in a column. Comparison of these populations enabled greater mechanistic understanding of the role of bacteria in the leaching of chalcopyrite. This work was the most comprehensive attempt to date made to delineate the complex microbiological/mineral actions using analysis of population dynamics from a mixed inoculum. It was found that the iron oxidiser Thiobacillus ferrooxidans dominated within the columns and leach solutions. The sulphur oxidiser Sulfobacillus thermosulfidooxidans was also prevalent in the columns, particularly during the period of rapid chalcopyrite oxidation. The high temperature, ferric leaching of chalcopyrite was unexpectedly poor in the first round of large columns. The reason for the low extraction was attributed to an increase in pH down the column, resulting in excessive goethite precipitation. The solution flowrate (velocity) was increased by ten times in subsequent columns. There were no operational problems (e.g. break-up of ore agglomerates). The increase in flowrate resulted in a high yield of copper. The kinetics of extraction were faster than a corresponding bacterial leach, confirming the potential advantage of a high temperature leach. The small column studies highlighted that it was important to get an even distribution of silver down the stope to enable maximum catalytic effect. If the ore were agglomerated, silver would be added with acid at that point. However, it may not always be possible to agglomerate the ore. For example, the process may be used in-situ on a fractured ore body, or on an ore that has a low fines content, and hence does not require agglomeration. Various complexing agents were tested for their ability to distribute silver at the start of the leach and to recover silver at the end of the leach. For instance when silver was complexed with thiourea and then trickled through the ore, an even distribution of silver was achieved. After leaching was completed, a thiourea wash recovered a significant amount of the silver. These two techniques minimised the amount of silver required and thus significantly added to the economic viability of the process. The success of the technical work has led to an evaluation of the process in the field. A flowsheet was developed for the high temperature, in-stope ferric leach of chalcopyrite. An economic analysis was performed that illustrated the process would be viable in certain situations. An engineering study considered issues such as acid consumption, aeration, silver distribution, silver recovery and a heat balance of the stope.

Supported stope mining is one of the most common types of mining in the modern day gold mining industry. The excavated regions, where ore is extracted, are supported with a combination of roof-bolting, timber packs, backfill, timber props and mechanical prop technologies. In order to install a support system that will be able to absorb the energy released by the elastic movement of the surrounding rock mass and support the unstable hanging wall, it is necessary for the rock engineer to know how the individual types of support will react to different load conditions in order to design a safe support system. Current support systems are developed using knowledge from past experience and trial and error processes. These are expensive and time consuming methods that can possibly be improved and made more cost effective by using modern design techniques. A study was conducted to determine the feasibility of the application of Finite Element Modelling (FEM) to the deformation of a modern support unit under specified quasi-static and dynamic stope load conditions with the view to assist in the prediction of the operational performance of support units that cannot be experimentally tested due to a lack of test equipment, capabilities and facilities. The study was extended by investigating the theoretical possibility of buckling due to an impact load on the prop and the performance of the prop. To achieve this, a simulation was carried out using ANSYS™ transient structural software to determine whether it is possible to simulate the performance curve of a prop. Computerised methods were used to determine the possibility of failure due to buckling and the implications of buckling, if it occurs, on the performance of a specific support prop design. In summary this study proved that it is possible to simulate the performance curve of a friction prop design in order to compare the result obtained with the required performance, provided that the correct friction coefficients between prop mating surfaces are known. It also presents a methodology to investigate the theoretical effect of high velocity impact load on the buckling potential of a friction prop design and slender columns in general, which is highly applicable to these types of support. The methodologies used in this study can be applied to different designs of friction props, and possibly reduce the development costs and implementation time of these types of support units. / MIng (Mechanical Engineering), North-West University, Potchefstroom Campus, 2014