Modelling the optimum interface between open pit and underground mining for gold mines

The open pit to underground transition problem involves the decision of when, how and at what depth to transition from open pit (OP) to underground (UG). However, the current criteria guiding the process of the OP – UG transition are not well defined and documented as most mines rely on their project feasibility teams’ experiences. In addition, the methodologies used to address this problem have been based on deterministic approaches. The deterministic approaches cannot address the practicalities that mining companies face during decision-making, such as uncertainties in the geological models and optimisation parameters, thus rendering deterministic solutions inadequate. In order to address these shortcomings, this research reviewed the OP – UG transition problem from a stochastic or probabilistic perspective. To address the uncertainties in the geological models, simulated models were generated and used. In this study, transition indicators used for the OP - UG transition were Net Present Value (NPV), ratio of price to cost per ounce of gold, stripping ratio, processed ounces and average grade at the run of mine pad. These indicators were used to compare four individual case study mines; with AngloGold Ashanti’s Sunrise Dam Gold Mine in Australia, which made the OP – UG transition in 2004 and hence develop an OP – UG transition model. Sunrise Dam Gold Mine is a suitable mine for providing baseline values because it recently made the OP-UG transition. Only four case study mines were used because it took nine months to generate transition indicators for each case study mine. A generic model was developed from the results of the four case studies to help mining companies make the OP - UG transition decision. The model uses a set of transition indicators that trigger the decision while recognising the uncertainties in the geological models, future mineral price as well as cost and processing parameters. From the generic model, mines can transition when the margin (gold price to cost per ounce ratio) is greater than 2.0; grade is between 4 g/t and 9 g/t, stripping ratio between 3 and 15 m3/t and positive NPV depending on the type of deposit. With this model mines can now transition when the critical conditions of the transition indicators (gold price to cost per ounce, grade and stripping ratio) are achieved. The model also uses the set of transition indicators to model the probabilistic nature of the OP-UG interface. The derived generic model will help mining companies in their annual reviews to assess the OP - UG interface and make decisions early enough with regard to transition timing

Evaluation of methods for stope design in mining and potential of improvement by pre-investigations

The importance of stope design for mine planning is considerable. Therefore, stope design and its challenges have been in the focus of research for the past 50 years. Empirical, numerical and analytical methods for stope design have been developed over the past decades in order to improve this process. This thesis is assessing which areas for improvement there still are and which problems are still only unsatisfactorily solved. After establishing background knowledge about the importance of stope design for mine planning and evaluating the factors influencing stope design, the focus is laid on the development of stope design methods in the past, as well as current research related to the topic, to create a comprehensive overview of recent and future developments. This is done by means of a literature review and research analysis. On the other side, the mining industry´s needs and challenges related to stope design are assessed, by means of survey, mine visit and interview. The insights gained in both parts are compared and checked for potential harmonies and disharmonies. Finally, from those conclusions practical recommendations for the GAGS-project are extracted and consecutively presented. In stope design research the focus and dominance of empirical methods has slowly shifted towards more research being conducted in the area of numerical and analytical methods. It can also be concluded that numerical methods and personal expertise are far more important for stope design within industry than commonly assumed. It was identified that in order to improve stope design, it is desired to increase the amount of geotechnical data acquired, the software improved, and stope design integrated within the general mine planning process. Additionally, interesting insights were gained by an in-depth analysis of survey responses, for example, the outstanding importance of the cut-off grade for stope design within gold mining operations. In order to allow for an optimal acceptance of novel geotechnical methods for stope design, the acquired data should be implementable into stope design within three days, preferably be compatible or implemented within a software and allow for stope design to be integrated into general mine planning. To promote the benefits a comprehensive scientific case-study demonstrating the realized benefits should be performed

Cut-off grade optimisation for a bimetallic deposit: case study of the Ruashi Mine Copper-Cobalt deposit

A research report submitted to the Faculty of Engineering and the Built Environment, University of the Witwatersrand, Johannesburg, in partial fulfilment of the requirements for the degree of Master of Science in Engineering. Johannesburg, 2017The research was driven by the need to optimise the Ruashi Mining operation to prevent further high-grading without destroying the value of the mine. Ruashi Mining incurred a five-year stripping backlog caused by the drive to reduce costs. As a result of this, a decline in metal production was imminent in the subsequent years. The study was conducted mainly using SimSched Direct Block Scheduler (SimSched DBS) in comparison with schedules from Datamine Net Present Value Scheduler (NPVS) and MineSched. The scenarios investigated have shown that running a mine based on break-even cut-off grade does not optimise the net present value of an operation as shown through the results of Ruashi Mining. The research also proved that royalty affects the cut-off grade for Ruashi Mining, and cannot be ignored. The proper scheduling of wasting stripping resulting from cut-off grade optimisation has contributed to a timeous exposure of high grade to avert the decline in metal production. All the three scenarios have shown that high grade ore can be availed on time, thus producing a smooth metal output for the life of mine. Cut-off grade optimisation is very crucial for any mining organisation as it is the main driver of value. Ore reserves are important in the determination of a company’s share price. High cut-off grade results in fewer reserves, and vice versa. Since mineral reserves are the source of revenue, therefore, the higher the reserves, the higher the revenue. Low cut-off grade may result in the processing of material that does not give high profit at the beginning of the life of mine. This, therefore, lowers the mining company’s net present value. This makes it imperative to optimise the cut-off grade during the mine life in order to optimise the net present value. During mining operations, there are various stakeholders whose interests must be considered during cut-off grade optimisation because they derive many benefits from the mine. These stakeholders include shareholders, employees, government, the community and non-governmental organisations. Cut-off grade optimisation has shown that there is an opportunity to improve the net present value of Ruashi Mining. SimSched gives a higher net present value (NPV) compared to the current Ruashi life of mine schedule. This indicates that SimSched can be used to improve the NPV for Ruashi by producing an optimised schedule. It is important to note though that there is need for the software to have provisions to take into account the initial stockpile status so that there is a holistic approach to the schedule optimisation. The grade-tonnage curve is steeper closer to zero implying that a small change in cut-off grade has a huge impact on reserves. Based on the results of the study it was clear that optimisation in SimSched DBS results in a steeply declining cut-off grade policy compared to NPVS. In addition, optimisation in SimSched leads to highly accelerated mining rate and massive stockpiling. Royalty is a cost which has to be incorporated in cut-off grade optimisation. The study has shown that the cut-off grade for Ruashi is increased by 19.8%. Ignoring royalty may result in overvaluing of an operation. Environmental considerations favour the optimisation of the use of the mieral resources. Consideration of environmental costs lowered the cut-off grade for Ruashi by 16%.MT 201

Optimization of blasting parameters in opencast mines

Drilling and blasting are the major unit operations in opencast mining. Inspite of the best efforts to introduce mechanization in the opencast mines, blasting continue to dominate the production. Therefore to cut down the cost of production optimal fragmentation from properly designed blasting pattern has to be achieved. Proper adoption of drilling and blasting can contribute significantly towards profitability and therefore optimization of these parameters is essential. Introduction Rock breaking by drilling and blasting is the first phase of the production cycle in most of the mining operations. Optimization of this operation is very important as the fragmentation obtained thereby affects the cost of the entire gamut of interrelated mining activities, such as drilling, blasting, loading, hauling, crushing and to some extent grinding. Optimization of rock breaking by drilling and blasting is sometimes understood to mean minimum cost in the implementation of these two individual operations. However, a minimum cost for breaking rock may not be in the best interest of the overall mining system. A little more money spent in the rock-breaking operation can be recovered later from the system and the aim of the coordinator of the mining work should be to achieve a minimum combined cost of drilling, blasting, loading, hauling, crushing and grinding. Only a “balance sheet” of total cost of the full gamut of mining operations vis-à-vis production achieved can establish whether the very first phase- rock breaking- was “optimum” financially; leaving aside factors of human safety. An optimum blast is also associated with the most efficient utilization of blasting energy in the rock- breaking process, reducing blasting cost through less explosive consumption and less wastage of explosive energy in blasting, less throw of materials, and reduction of blast vibration resulting in greater degrees of safety and stability to the nearby structures. Development of a Blast Optimization Model Selection of proper explosive in any blasting round is an important aspect of optimum blast design. Basic parameters include VOD of explosive (m/s), Density (g/cc), Characteristic impedance, Energy output (cal/gm), and Explosive type (ANFO, Slurry, Emulsion etc.). However, all these parameters can not be taken for optimizing the blasting method successfully. Some of the parameters are taken for minimizing the blasting cost. These cost reduction and optimum blast design parameter will give an economical result. The parameters are i. Drill hole diameter, ii. Powder factor (desired), iii. Cost of explosive, iv. Numbers of holes required to blast. Methodology The study of the various parameters of blasting suggests that the powder factor should be constant as per the requirement. The number of holes desired as per the explosive, the drill ihole diameter as available and the cost of explosive are kept as input. The spacing, bench height, burden, charge per hole as depending on the previous parameters can be calculated. From the different input and calculated parameters the total cost of the method is calculated and the least expensive method is selected as the optimized model. Blasting related information were collected from three different mines of Mahanadi Coalfields Ltd.(MCL) for implementation of the optimization model. A program was designed using visual basic on .net platform taking the above parameters into consideration to select the optimized model. It was observed that the program gives satisfactory results. A sample output of the program is as presented below: Conclusion The blast optimization model has been developed with simple methodologies which can be adopted by the mining industry to compare the explosive costs and achieve better blasting results and. The model developed is a user friendly one, since by keeping the powder factor and number of choices of explosives available as constant and by varying the parameters like drill hole diameter, number of holes and cost of explosives one can compare the explosive performance and accordingly take a decision to select the proper type of explosives for blasting. It may be noted, that the model has been developed based on case studies of three different mines of MCL, and it can be modified with collection of information from a large number of mines. References Nanda, N.K. (2003), “Optimization of mine production system through operation research techniques”, 19 th World Mining Congress, New Delhi, November, pp.583-595. Pal Roy, P. (2005), “Terms and parameters influencing mine and ground excavations”, Rock blasting effects and operations, pp. 17-22

Integrating materials supply in strategic mine planning of underground coal mines

In July 2005 the Australian Coal Industry’s Research Program (ACARP) commissioned Gary Gibson to identify constraints that would prevent development production rates from achieving full capacity. A “TOP 5” constraint was “The logistics of supply transport distribution and handling of roof support consumables is an issue at older extensive mines immediately while the achievement of higher development rates will compound this issue at most mines.” Then in 2020, Walker, Harvey, Baafi, Kiridena, and Porter were commissioned by ACARP to investigate Australian best practice and progress made since Gibson’s 2005 report. This report was titled: - “Benchmarking study in underground coal mining logistics.” It found that even though logistics continue to be recognised as a critical constraint across many operations particularly at a tactical / day to day level, no strategic thought had been given to logistics in underground coal mines, rather it was always assumed that logistics could keep up with any future planned design and productivity. This subsequently meant that without estimating the impact of any logistical constraint in a life of mine plan, the risk of overvaluing a mining operation is high. This thesis attempts to rectify this shortfall and has developed a system to strategically identify logistics bottlenecks and the impacts that mine planning parameters might have on these at any point in time throughout a life of mine plan. By identifying any logistics constraints as early as possible, the best opportunity to rectify the problem at the least expense is realised. At the very worst if a logistics constraint was unsolvable then it could be understood, planned for, and reflected in the mine’s ongoing financial valuations. The system developed in this thesis, using a suite of unique algorithms, is designed to “bolt onto” existing mine plans in the XPAC mine scheduling software package, and identify at a strategic level the number of material delivery loads required to maintain planned productivity for a mining operation. Once an event was identified the system then drills down using FlexSim discrete event simulation to a tactical level to confirm the predicted impact and understand if a solution can be transferred back as a long-term solution. Most importantly the system developed in this thesis was designed to communicate to multiple non-technical stakeholders through simple graphical outputs if there is a risk to planned production levels due to a logistics constraint

PROCJENA STABILNOSTI KOSINA U VLAŽNOJ TROPSKOJ REGIJI: STUDIJA SLUČAJA RUDNIKA UGLJENA U JUŽNOM KALIMANTANU, INDONEZIJA

PT.X, a coal mining company in South Kalimantan, Indonesia, plans to use the highwall mining method to excavate marginal reserves on the final slope to maintain production. However, the stability of the slope and determination of the highwall mining dimensions are major concerns due to unfavourable rock mass conditions caused by intensive weathering and tectonics. This paper aims to evaluate the feasibility of highwall mining in the study area using empirical, analytical and numerical methods. The innovation of this research is the integration of these methods, which include rock mass classification, analytical calculation of load and rock support strength, 2D and 3D numerical modelling, and estimation of recovered coal from the highwall design. The initial condition assessment using rock mass classification and analytics calculation of the mining geometry model determined mine openings and pillar dimensions. Numerical modelling re-evaluated the geometry models to obtain an optimal design. The suggested optimal thickness, mine opening, web pillars, and barrier pillars are 3.20, 3.00, 3.50, and 4.00 m, respectively, with four web pillars in one panel at Seam-C and 2.50, 3.00, 3.50, and 4.00 m with four web pillars in one panel at Seam-D. The recovery of coal for Seam-C and Seam-D is estimated to be 40.54%. Deformation was found to have the closest relationship with the dimensions of the mine opening, and the safety factor is most sensitive to changes in the depth of the mine opening. This study provides a reference for future highwall mining in Indonesia and other regions with similar conditions.Kako bi unaprijedila proizvodnju, tvrtka za eksploataciju ugljena PT. X u Južnom Kalimantanu u Indoneziji, planira koristiti visokočelnu metodu iskopavanja preostalih rezervi ugljena iza završne kosine. Pri tome su glavni izazovi stabilnost i određivanje dimenzija otkopavanja zbog nepovoljnih uvjeta stijenske mase uzrokovanih intenzivnim vremenskim uvjetima i tektonikom. Cilj ovog rada je empirijskim, analitičkim i numeričkim metodama procijeniti izvedivost visokočelne metode eksploatacije u istraživanom području. Inovacija u ovom istraživanju je integracija raznih metoda, koje uključuju klasifikaciju stijenske mase, analitički proračun opterećenja i čvrstoće stijena, 2D i 3D numeričko modeliranje te procjenu dobivenog ugljena pri projektiranju visokog čela. Početno stanje je procijenjeno pomoću klasifikacije stijenske mase, a analitičkim proračunom u geometrijskom modelu utvrđene su širine iskopa i dimenzije zaštitnih stupova. Zatim je numeričkim modeliranjem ponovno procijenjena geometrija modela kako bi se dobio optimalan dizajn. Predložena je optimalna debljina i širina iskopa te raspored zaštitnih stupova od 3,20, 3,00, 3,50 odnosno 4,00 m, s četiri raspoređena stupa u jednom panelnom otkopu za sloj C te 2,50, 3,00, 3,50 i 4,00 m s četiri raspoređena stupa u jednom panelnom otkopu za sloj D. Procjena iskorištenje ugljena za sloj C i D je 40,54%. Utvrđena je međuovisnost deformacija s dimenzijama otkpnog usjeka, a faktor sigurnosti je najpromjenjiviji kod promjene dubine iskopa. Ova studija pruža preporuke za buduće eksploatacije sa visokočelnom metodom u Indoneziji i drugim regijama sa sličnim uvjetima