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Acid mine drainage (AMD) is one of the biggest environmental issues faced by the mining industry. The concerns towards AMD rises when the waste rock and tailings are enriched in pyrite and/or pyrrhotite. The presence of trace of heavy metals bearing minerals increases the environmental hazard of such material. And the large volume of solid waste produced during operation leads to a high pollution risk. AMD is characterized by the release of acidic runoff water enriched in sulfate and metals. However, the impacts associated with AMD are beyond environmental, besides disturbing aquatic life and contaminating soil and water streams, the areas affected by AMD faces the visual impact associated with its characteristic “ochre” color and can have their water supply or tourism affected by the pollution. Additionally, the expenditures to recover these sites are extremely high. A reliable AMD assessment is crucial to design an appropriate tailings management strategy capable of preventing AMD generation and a suitable rehabilitation plan able to reduce the footprint of such operations in the surrounding areas. The oxidation of iron-bearing sulfides (pyrite, pyrrhotite and marcasite) occurs naturally when in contact with moister and oxygen, however, mining increases the exposure of these rocks whereas the comminution process increases the reactivity of the minerals by affecting its degree of liberation and particle size. Hence, the oxidation process is enhanced by mining operations. Acid mine drainage generation depends mainly on the ability of neutralizing minerals to buffer the acidity produced by pyrite oxidation. Besides that, particle size and liberation degree define the reactivity of the acidproducing and acid-neutralizing phases. Hence, these parameters influence the oxidation rate and the overall geochemical evolution of the leachate released. The traditional methods used to assess AMD does not consider particle size and liberation degree. However, to achieve a reliable AMD prediction, account to these parameters are of utmost importance. This master thesis aims to analyze the effect of liberation degree in AMD generation using tailings coming from Dundee Precious Metals Chelopech (DPM-Ch). The tailings produced by the concentration process in DPM-Ch is characterized by enrichment in pyrite. Quartz and kaolinite are the main non-sulfide gangue whereas no carbonate mineral is present in the ore. Thus, no neutralizing mineral is present in the tailing assemblage. Moreover, the presence of trace arsenic bearing minerals which are not recovered during the process increases the environmental concern toward this material. Particularly because this element is mobile under acidic conditions and can be toxic to humans, plants and animals in low concentration. To achieve this objective an integrated approach, that combined process mineralogy techniques and geochemical testwork (static and kinetic test), was used to assess the AMD potential of the tailings from DPM-Ch processing plant. Samples from the tailings management facility (TMF) were collected and pyrite flotation was performed aiming to recover the free pyrite while the locked pyrite remained in the tailings. Thereafter, SEM-EDS analysis was done to characterize each flotation stream regarding its modal mineralogy, liberation degree and particle size. Geochemical tests (static and kinetic tests) were performed in the pre-floated samples and tailings after pyrite flotation. The results showed that the reactivity of the tailings is highly dependent on the liberation degree and that the finer fraction dictates the overall geochemical behavior of the material. Furthermore, 40% of the tailings produced in Chelopech is used to backfill the stopes. This technique enhances the productivity of the mine, increases the safety of the underground workplace and decreases the volume of tailings being handled on the surface. In addition to that, the cemented paste has the potential of decreasing pyrite reactivity by encapsulating this mineral inside a cemented matrix. However, a proper study of the geochemical behavior of the paste should be conducted to evaluate the environmental performance of the cemented paste backfill (CPB). This was done by using monolithic leaching test (MLT) and mineralogical analysis. The results showed that the calcium phases present in the cement buffer the acidic water that percolates the stopes. Likewise, this technique may reduce pyrite oxidation.
Acid mine drainage --- Tailing management facility --- Static test --- Kinetic test --- Liberation degree --- Sulfide minerals --- mineralogy --- cemented paste backfill --- environmental behavior --- tailings --- geometallurgy --- monolithic leaching test --- Ingénierie, informatique & technologie > Géologie, ingénierie du pétrole & des mines
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In this project, the ore of the project of expansion of Draa Sfar mine is studied. The deposit is a Cu-Pb-Zn polymetallic Volcanogenic Massive Sulphide. The project extension contains three main ore-types: the polymetallic mineralisation from -1200 m to -1300 m, the polymetallic mineralisation from -1300 m to -1500 m and a copper-rich polymetallic ore. Targeted minerals are galena for lead, chalcopyrite for copper and sphalerite for zinc. The ores of the extension are first studied to determine its quantitative mineralogy. The ores of the deposit extension are compared to the ore that is presently exploited in Draa Sfar. Then, in a second stage, their behaviour is estimated in a laboratory flowsheet similar to the one of the the flotation plant used to process the ore currently extracted from Draa Sfar mine called reference ore. The flowsheet comprises multiple grinding stages in between a lead, a copper and finally a zinc flotation circuits. To fulfil both objectives, the three ores of the extension project and the present ore are sampled and analysed. Then, they are floated in laboratory tests and their products are analysed on bulk and sized samples, SEM is also used. A sampling of the plant is also conducted, but on a too short period to provide reliable results. Ores of the extension project appear poorer in lead and zinc than reference ore, they are also richer in copper. In the lower part of the deposit, -1200 m, the liberation degree of galena is decreasing and the one of chalcopyrite increasing compared to the reference ore. These characteristics of the future feed of the plant suggest to modify the flowsheet to a copper flotation followed by lead and zinc flotations. At laboratory scale, the reference ore produces lead, copper and zinc concentrates with higher recoveries than in the ores of the extension project. In the lead, copper and zinc circuit, the concentrates are diluted by an entrainment of sulphide and silicate gangue pointing to a lack of selectivity of the three flotation circuits. In the lead circuit, the low recoveries in the concentrates of the ores of the extension project compared to reference ore can be related to a lack of liberation of galena leading to losses to the middling streams of the copper circuit. In the copper concentrate, copper grade is higher in ores from the extension thanks to the coarser liberation of chalcopyrite in the reference.
Cu-Pb-Zn polymetallic ores, quantitative mineralogy, process mineralogy, geometallurgy, flotation, laboratory test, laboratory scale, plant scale, sampling, masse balance, size by size study, SEM, liberation degree, mineral associations, sulphide gangue --- Ingénierie, informatique & technologie > Géologie, ingénierie du pétrole & des mines
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