Surface mining operations can have a significant visual impact on the surrounding landscape and communities. There have been numerous cases of local populations pushing back against the development of mining operations near their communities. This study aims to determine whether it is possible to assess the visual impact of surface mining operations before the development has started. With this assessment it would become possible to make informed decisions as to pitshell development and selection in terms of visual impact. A tool was developed to determine whether it is possible to make this assessment. This tool must be able to be used to determine the visual impact of two or more scenarios in the pre-feasibility phase of development to assist in the decision of
which scenario will be developed further in the feasibility phase. In this context, the visual impact is defined as the perceived change in the landscape as a result of mine development on local communities around the surface mine development. The scenarios are defined as the option of pitshells designed during the pre-feasibility phase of mine development.

The tool makes a distinction between the physical changes in the landscape caused by the surface mining operation and how this physical change is perceived by anyone viewing the operation.

The tool is GIS based and utilizes the free open-source software QGIS to calculate the physical change in the landscape. The tool calculates the vertical and horizontal visibility angle. The visibility angle describes the extent that the change in the landscape takes up in the view of an observer. In addition
to that, the tool calculates the contrast between the changed colour in the landscape and its surroundings, which describes how much the change stands out. It requires two sets of data to work: a digital elevation model of the prospected surface mine pitshell and the surrounding landscape and a
RGB satellite image of the surrounding area. The DEM provides elevation data for every square metre of the landscape to the GIS, while the satellite image provides RGB colour data. All other data can be generated from the previously mentioned data features.

The tool is tested using a case study, in which several distinct pitshell scenarios are compared against one another on their visual impact. The case study is based on the extension of a limestone quarry in
Belgium, which aims to secure reserve for future operations. The area of interest is surrounded by several small to medium size villages, which would be affected visually by the extension. Therefore, a comparison should be made regarding the visual impact of the several pitshell scenarios. The results showed it is possible to determine the visual impact of surface mining operations before the
development has started. On this basis, it is possible to compare different pitshell scenarios and determine which scenario would be least impactful in terms of visual impact on the local community and surrounding landscape. The reliability of the tool can only be assessed theoretically, however a high level of accuracy is achieved by adhering to guiding principles which are set out in this thesis.

Open-pit mining blast patterns are designed to safely break rock mass at desired boundaries and to produce uniformly sized fragments that lead to the optimal performance of subsequent mining operations. An issue that is often overlooked, however, is the deviation of drillhole locations with respect to their coordinates as designed in the blasting plan. This causes changes in the spatial distribution of explosive energy; some areas of the blast may receive a surplus of explosive energy, while others see a reduction. This results in less uniform fragmentation, producing more fines and increasing chances of oversized boulders, which reduces the efficiency of loading, hauling, and crushing operations. Additionally, deviated drillholes can lead to slope stability issues, flyrock, and uneven benches. To mitigate these problems, this study attempts to reduce differences in the explosive energy distribution (EED) through an optimization approach that adjusts the height of explosive columns in blastholes. A case study combines planned drillhole coordinates of two blast patterns with both simulated deviations and actual deviated drillhole locations. The corresponding benches are discretized into block models, for which the explosive energy distribution is calculated from the planned drillhole configuration as well as the deviated holes. The optimization objective is then to minimize block-to-block differences in explosive energy by making discrete adjustments to the height of individual column charges. Since mathematical programming is not possible due to the complexity of the evaluated formula, solutions are generated by metaheuristic algorithms. Among the Genetic Algorithm (GA) and Tabu Search (TS), the latter is preferred for its local search neighborhood that gives superior computation times. An inconvenience of the block-to-block comparison is the introduction of very large differences in explosive energy in blocks nearest to deviated blastholes, which cannot be accounted for by any charge adjustments except reductions in the deviated hole itself. Because optimization should focus on manageable differences in lower-valued blocks, these high-valued blocks are discarded by applying a cap based on data percentiles. Obtained results show that optimized charging adjustments vary with the chosen percentile, which makes it challenging to define comprehensive truly optimal solutions. A bias for charge reductions at high percentiles gradually shifts towards a bias for charge increases at lower values. Because a bias-induced imbalance in the total amount of explosives (compared to the original charging) is undesirable, the setting closest to equilibrium is selected, which is the use of only the lower 50\% of data. In this configuration, optimization of the case study blast patterns results in improvements of the objective value of 0.53-1.54\%, or 2.14-3.94\% when only the blocks affected by the recommended charge adjustments are considered. The block-to-block comparison handles any interchangement of block values as unwanted differences, which should not always be the case. Therefore, the true improvements to the explosive energy distribution might be better expressed using an alternative optimization objective. Computation times could be further reduced by expanding on the developed algorithm, and combination with prior research would be helpful to include any GPS-related inaccuracies in the actual drillhole locations.

To optimise mining in a safe, geotechnically controlled and economical most beneficial way in Mine 2 of the Stjernøy mine, a choice has been made to review the current mining method, sublevel open stoping, and a different mining method, vertical crater retreat. The Stjernøy mine is an open pit and underground mine for Nepheline Syenite and this thesis focuses on Mine 2, the underground mine which is partially completed. The stopes of Mine 2 are closely located to old stopes, which can cause stability issues. Between the new and old stopes is a barrier pillar which is a key element in the stability of the underground mine and the infrastructure for the open pit operation as well. This project consists of field work and laboratory testing to create a basis for the numerical stability analysis of the mine and the creation of a cost model for the two different mining methods. The results of the cost modeling show that the use of sublevel open stoping will result in a cost per tonne of 9.86 NOK and the use of vertical crater retreat will result in a cost per tonne of 10.39 NOK. The use of sublevel open stoping as mining method is the most economically beneficial option for the mining operation. It is less expensive as a mining method and it is a proven method at this mine site. The vertical crater retreat method uses a more expensive blasting method, with the same burden and spacing parameters as the sublevel open stoping. The results of the field work and the laboratory testing provided important parameters for the numerical modelling, such as the strength of the material. Based on these results the numerical modelling also showed that it is beneficial for the stability of the pillar if some stopes near the barrier pillar are not extracted. The extraction of these stopes will decrease the factor of safety of certain areas of the barrier pillar to one or for some areas below one. The stope size has also been reviewed. To make sure that the stopes are well connected to drawpoints, it is important to keep them the same size, which is a stable stope size for the roof. Smaller stope sizes are found to be a bit more stable, however the effect of the increasing amount of pillars is not known on the overall stability.

The aim of this master thesis is to improve a raw materials certification scheme by checking and validating the application matrix and the coherent hazards developed for the Certification of Raw materials (CERA) project. The scope of the original application matrix included exploration, mining, physical processing, chemical processing, smelting and refining. The validation in this thesis was focussed on the iron ore value chain. This validation showed that important methods associated with these processes were missing in the application matrix. Two main sections that are absent are recycling and storage and transportation, which have been added to the application matrix including the main methods used within those two sections. The risks and hazards developed by CERA, associated with the application matrix’s processes and methods, are transformed to only hazards. Next to this, new hazards are suggested to make the hazards list more holistically applicable. After the addition of recycling to the application matrix, an analysis on the addition of recycling to the CERA Performance Standard (CPS) and to the CERA Chain of Custody Standard (CCS) was done. This analysis showed that adding recycling to the Performance Standard deals with few challenges due to it being an industrial activity with many similarities to methods used in processing and smelting. Therefore, with the additions that are proposed in the validation of the matrix and the hazards analysis, recycling can be easily added to the CPS. This is done by including collection centres, recycling facilities and scrap transport in the standard under the general term ‘recycling facilities’. Adding recycled material to the CCS can be done in different ways, this thesis proposes multiple options of which three are viable. First, three types of recyclable materials have been distinguished: Recyclables of which the origin is known (1); Recyclables of which the origin is unknown and no further information is available (2); Recyclables of which the origin is unknown but a due diligence can show that it has sufficient added value to the raw materials value chain with regard to responsibility and sustainability (3). In all cases, CERA should include the certification of recycled material of which the origin is known. The first viable option is to only certify type one with the original CCS certificate and exclude the other two types from the scope. This requires no extra certification or requirements by CERA. The second viable option for CERA would be to also certify type three within the original CCS certification standard, and add a ‘recycled label’ to materials of type two. The third viable option would be for CERA to develop a ‘CCS recycled’ certificate for materials of the third type mentioned, and add a ‘recycled label’ to materials of type two. CERA should decide which option is best according to their basic criteria and values. If CERA decides to select option one, CERA safeguards their integrity but the option is lacking holistic applicability. If option two is preferred, CERA increases their holistic applicability level, however, the integrity of CERA might be at stake. If CERA decides to select option three, holistic applicability is maximized and their integrity is safeguarded, however, the system of CERA is complicated due to different levels of certifying.

Current evacuation plans for underground mines are inefficient and outdated. Simulations have proven that smart evacuation can significantly increase the efficiency of escape times in case of an emergency. In a smart evacuation, miners are given real time instructions on a smart device (such as a smart-watch) about their route to a safe haven if an emergency occurs. In order to make smart evacuation possible, an optimization model is needed that determines the most efficient route to safety. Currently, shortest path algorithms, such as Dijkstra’s algorithm, Floyd-Warshall’s algorithm, or ant colony optimization are used for solving the optimization models. These algorithms, however, are computationally inefficient (Dijkstra and Floyd-Warshall) or may not provide the optimal solution (ant colony optimization). In this thesis, a mathematical programming method will be used to calculate the most efficient escape plan. Mathematical programming translates the goal of finding the most efficient escape solution to an optimization problem. The goal of this optimization problem is to minimize the total distance travelled by all individuals that are underground in case of an emergency. This is done by setting an objective function, which sums up all the distances travelled by the miners, and a set of constraints, which are used to localize the miners and safe havens. The ensuing problem statement is solved using the network simplex algorithm. This principle was used to calculate escape solutions in four different scenarios: with or without blocked pathways and with or without correcting the distances for the stamina of the individual miner. Four different types of results were generated: the division of miners among the safe havens, the total distance travelled by all individuals taking part in the evacuation, the path of an individual miner (called Miner X), and the running times of the different scenarios. It was found that blocked paths can have a large impact on the division of miners among safe havens and can significantly increase the total distance travelled. Using stamina categories increases the path length of Miner X (who has a high stamina) and also the running times of the algorithm. It can be concluded that the algorithm works for all four scenarios. As including blocked paths does not give a time penalty, but will locate trapped miners and send their colleagues on safer escape routes, using this feature has no downsides. The usefulness of adding stamina categories to the algorithm can be debated. The running time of the algorithm increases, while the solutions for realistic numbers of miners do not change. Furthermore, there are philosophical and social questions about the ethics of using these types of categories.

Large quantities of mine tailings are stored in mine tailings storage facilities. These facilities are only costing money for mining companies. In addition, these tailings storage facilities are prone to failure. To prevent this from happening these facilities have to be reinforced. This research introduces a concept to use geotextile tubes filled with mine tailings as reinforcing construction elements. That way value is assigned to the mine tailings while also making the storage facilities safer. However, pumping mine tailings in geotextile tubes is not an option due to the small particle sizes which will flow through the openings in the geotextile fabric. In addition, the contaminations in the mine tailings will also be mobilised using this method. Therefore, this research investigates potential enhancement of mine tailings to make them suitable to be pumped in geotextile tubes. The potential enhancement is in the form of flocculation and remediation. While flocculation increases the dewatering properties of mine tailings, the remediation will prevent the contaminations from mobilising. Test performed in this research include a series of jar tests, cone tests, two different strength tests and hanging bag tests. The results show that when flocculants are added to the tailings, the dewatering rate increases. The flocculant type which proved the most effective for these specific tailings type is a flocculant with an average molecular weight and average anionic value. Adding coagulant to the tailings further increases the dewatering rate of the tailings. When Ordinary Portland Cement is added, to stabilise and solidify the tailings, the dewatering rate increases even more. In addition, adding Ordinary Portland Cement to the tailings resulted in an increase of shear strength and unconfined compressive strength. After seven days of curing, the highest strength of curing was measured for the mixtures where fifteen mass percentage Ordinary Portland Cement was added. The mixture where the highest strength and the fastest dewatering rate was measured, was the mixture composes of mine tailings, flocculants, coagulants and fifteen mass percentage Ordinary Portland Cement.
With these results, it may be concluded that mine tailings can be used to fill geotextile tubes to form reinforcing construction elements if its dewatering behaviour is enhanced and its contaminants remediated.

Repetitive construction projects allow for planning to be done systematically, but performance is hard to predict due to inherent uncertainties. The goal of this research is to evaluate planning choices in repetitive construction projects. In order to evaluate the planning of these projects to quantify the risk of overruns, an adaptable digitized model is needed which can run realistic probabilistic simulations. The model was supplemented with realistic effects such as probabilistic distributions and learning curves to represent the inherent uncertainties in the progression of the project. The model is able to estimate the total duration of a project, the time-dependent costs and resource utilization percentage. Even more so, the model was used to give an estimation of the risk of overruns as a probability. The model has been used to evaluate decisions that can be taken when planning a project. To test how these planning decisions affect the project outcomes, they were tested in a test case set-up. The influence of one planning decision was increased and the outcomes measured. These planning decisions consist of 1) the use of multiple resources crews per process, 2) the use of batching, 3) using time buffers and 4) multiskilled crews. For decreasing the project duration, extra crews can be added. Time-dependent costs are decreased by adding buffers or working with larger batches. Improving the utilization percentage is done by using multiskilled crews and adding buffers. The results are relative and can be interpreted as rules of thumb, but are indicative of the power of planning decisions to steer project performance without altering the scope or the construction method. To show that the model indeed corresponds to these rules of thumb, they were applied in two projects, yielding promising results, both for total duration and time-dependent costs. Using the understanding of the effects of planning decisions, repetitive construction projects can be improved to increase productivity and decrease time-dependent costs significantly.

An investigation in the possibility to integrate fleet selection (through FPC) in the mine design process, allowing for the analysis of various potential haul road scenarios and from there being able to make a decision on an optimal fleet selection.

In the mining industry the initial capital expenditure budgeted in the feasibility study is being overrun structurally. These capital expenditure overruns have long been recognised, yet no method is available predicting the probability a mining project will experience an overrun.

This thesis introduces a method to assess the likelihood and magnitude of an initial capital expenditure overrun occurring for future gold mining projects. A database is populated with 65 gold mining projects that have experienced an overrun. For each project, 19 variables are described in the database that define the projects. The variables are analysed using individual variable analysis and multivariate regression analysis to investigate whether a correlation can be found between the variables and the initial capital expenditure overrun. Using the results of the individual variable analysis, a tool is developed to semi-quantify the risk a future gold mining project is exposed to regarding an initial capital expenditure overrun. The tool can visualise the risk a project is exposed to in a 2D risk matrix displaying the likelihood and magnitude of a possible overrun.

For 16 of the 19 variables a possible correlation is found. The variables are grouped together, representing root causes with regards to the initial capital expenditure overrun. The database analysis suggests that relatively small gold mining projects experience a larger capital expenditure overrun than large gold mining projects. In addition, projects owned by relatively large companies tend to experience a larger capital expenditure overrun. Furthermore, projects where a low number of payback years or a high internal rate of return presented in the bankable feasibility study indicate a profitable project tend to experience a smaller capital expenditure overrun. Finally, the database suggests a correlation between the relative gold price and the capital expenditure overrun. Projects from which the bankable feasibility study was published when the gold price was relatively high, tend to experience initial capital expenditure overruns more frequent and with a larger magnitude. Using the correlations resulting from the database analysis, a predictive tool is successfully developed for which future gold mining projects can be assessed on initial capital expenditure overrun risk.

Up until now, global mining companies have failed to create a fatality free environment for their employees, whether it be through increasingly difficult conditions or inadequate safety management approaches. Desiring to maintain the social license to operate, these companies are now looking into alternate ways of safety management with the aim of mitigating fatalities in the workplace.
In this study, a new risk management approach was applied to mitigate fatal incidents through the utilization of critical controls. The aim of this study was to create a scalable, minimally invasive proof-of-concept for AngloGold Ashanti that can successfully be implemented at any of the company’smining operations.
The foundation for this implementation was laid through a literature study blending contemporary safety and risk management approaches with Business Intelligence systems thinking and a perspective on industry best-practices. Subsequently, a framework of organizational requirements was set up based on 26 interviews with stakeholders in various departments within the company. Both design and implementation were conducted on-site at an AngloGold Ashanti operation in Western Australia.

The system was designed by adhering to organizational requirements, and ensuring that it is suitable to any mining environment. The designed Critical Control Management System was subsequently implemented at Sunrise Dam, one of AngloGold Ashanti’s Australian mining operations. To ensure that critical controls were also assessed at the operational level, a workplace inspection process was modified to generate control data. All sources of data subsequently were fed into a Business Intelligence environment enabling insight into critical control performance to all company stakeholders. Doing so informs decision-making on safety priorities company-wide, based on real-time data generated on the operational level.

Two case studies were performed to assess two of the most significant hazards at Sunrise Dam. The studies showed that the effectiveness of reactive controls changes irrespective of their compliance and performance. Furthermore, the influence of human factors within risk management remains difficult to quantify. Finally, it demonstrates the potential for integration of incident data into the Critical Control Management System, thus creating both leading and lagging indicators for safety performance.

The conclusion of this study is that an effective and scalable Critical Control Management System can be successfully implemented in a mining operation if the right conditions are generated. The approach of integration in existing processes demonstrates that companies can achieve greater control over fatality prevention without the need for an additional safety management system. On this basis, it is recommended that other operations are supported in creating an environment suitable for adaptation before Critical Control Management is implemented.