GEOLOGY, MINERALOGY AND CHEMISTRY OF THE GOLD ORE AND RELATED TAILINGS 

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GEOLOGY, MINERALOGY AND CHEMISTRY OF THE GOLD ORE AND RELATED TAILINGS

Gold was discovered in 1886 in the Witwatersrand and this region represents the largest known low-grade gold mineral deposit in the world (Adamson, 1973). The thickness of the Witwatersrand Supergroup reaches 7500 m and occupies an area of thousands of square kilometres in the Gauteng (former Transvaal) and the Free State Provinces. The production from the Witwatersrand Basin, since the first discovery, amounts to about 45 000 tons of gold (Au) and about 150 000 tons of uranium (Robb Meyer, 1995).
According to Feather and Koen (1975) the Witwatersrand conglomerate is a greyish, metamorphosed sedimentary rock consisting mainly of quartz (about 80 per cent), cemented by a fine-grained matrix of recrystallized quartz and associated with various phyllosillicates (i.e. a mixture of sericite and chlorite, and sometimes also pyrophyllite and chloritoid). The pebbles vary in composition, size, and colour but consist mainly of vein quartz. Round grains of pyrite, also known as buckshot pyrite, are often visible in the matrix, and sometimes are used as indicators for high gold concentrations. The gold is confined to the matrix of the conglomerates and is concentrated predominantly along the bedding planes of the conglomerate beds and on the footwall contact (Liebenberg, 1973).
According to Feather and Koen (1975) the Witwatersrand conglomerate is a greyish, metamorphosed sedimentary rock consisting mainly of quartz (about 80 per cent), cemented by a fine-grained matrix of recrystallized quartz and associated with various phyllosillicates (i.e. a mixture of sericite and chlorite, and sometimes also pyrophyllite and chloritoid). The pebbles vary in composition, size, and colour but consist mainly of vein quartz. Round grains of pyrite, also known as buckshot pyrite, are often visible in the matrix, and sometimes are used as indicators for high gold concentrations. The gold is confined to the matrix of the conglomerates and is concentrated predominantly along the bedding planes of the conglomerate beds and on the footwall contact (Liebenberg, 1973).
The conglomerates from the different mines are mineralogically very similar, but vary in the relative proportions of minerals comprising them. A detailed discussion on the mineralogy of the Witwatersrand Reefs is provided in Feather & Koen (1975).
It can be expected that the mineralogical composition of tailings from the recovery of gold can be derived from the gold ore. In this study, 16 tailings samples have been selected for the determination of mineral content by means of the semi-quantitative X-ray diffraction method (XRD). Table 2.3 presents the result of these analyses
reported high contents of phyrophyllite (max. 16 per cent) and sericite (max. 2 per cent) as well as quartz contents of 80-90 per cent in tailings material. Pyrophyllite and muscovite show similar physical properties to kaolinite: low expanding capabilities when hydrated (swelling and shrinking, with changes in moisture content), and a low cation exchange capacity.
It is important to note that these mineral phases in tailings impoundments control the pore water chemistry, thus affecting the chemical composition of acid mine drainage.
Gold has been also recovered in the East Rand area where the Black Reef Formation occurs at the base of the Transvaal Supergroup (Table 2.1), particularly where it is in close proximity to the underlying gold-bearing Witwatersrand beds (Liebenberg, 1973). In this context, Barton & Hallbauer (1996) reported average trace metal concentrations for pyrite grains of the Black Reef Formation of the Transvaal Supergroup and a summary is listed in Table 2.5.
According to the chemical composition of the tailings material, the high Sio2 values correspond to the high quartz content in the gold ore (Table 2.4). Carbonate occurs as traces in the Witwatersrand gold ore, but lime is added during the gold recovery process, resulting in an alkaline slurry and, thus in elevated carbonates contents in the tailings. However, these contents in the tailings are too low to provide sufficient acid neutralisation capacity to prevent the generation of the acid mine drainage.
The loss of ignition (LoI) as shown in Table 2.6a usually reflects the total content of organic matter and volatile elements such as CO2, H20, C, CI, F, Sand CN (cyanide). It is most unlikely that tailings contain any significant concentrations of organic
material as the content of kerogen is generally low in the gold ore. Cyanide is used amongst others during the gold recovery process to dissolve the gold and this process is described in more detail in the next paragraph. However, CN is unstable and decomposes rapidly if exposed to sunlight and the atmosphere (Adamson, 1973).
The following parameters may influence the trace element concentration in the mine tailings:
Fluctuations in the pyrite content of the mined ore; Dilution effect by the matrix; Metallurgical  separation  during the gold recovery process; Oxidation within the surface layer and migration into deeper zones of the impoundment.
The correlation coefficients of all measured elements (major and trace elements) were calculated and are presented in Table B-7 (Appendix B).
Subsequently, the mobility of various trace elements in 13 tailings samples was investigated. The extractable concentrations and the relevant threshold value for soils are presented in Table 5.1. In summary, high concentrations of sulphur (S) in the leachate are caused by the oxidation of sulphide minerals such as pyrite. Furthermore, Co, Cr, Cu, Ni, Pb and Zn exceed in the bulk of the leachate samples the soil standard used in this study.
The recovery of gold in South Africa is achieved by a number of mechanical and metallurgical processes (Adamson, 1973). The first step in the recovery is sorting to reduce the mass of ore milled by eliminating dilution with waste rock from run-of-mine feed. The waste rock is either deposited, or sold as construction material. The second step is crushing and milling to reduce the grain size of the ore to a size less than 0.1 mm. The fine-milled material is suspended in water and passed through a hydrocyclone in order to separate over-size material for recycling to the mill. In the next metallurgical step gold is recovered partly by gravity concentration (for coarse gold) and partly by cyanidation (for fine gold), where mine water from tailings ponds is used to augment the water in the milling circuit.
Various processes (not necessarily listed in the proper sequence) achieve the metallurgical recovery of gold:

• Gravity concentration;
• Thickening;
• Cyanidation;
• Filtration;
• Precipitation  and smelting;
• Carbon-in-pulp (CIP) process.

The fine material from the top outlet of the hydrocyclone, or overflow, has to be thickened by adding lime and organic compounds as flocculants (Funke, 1990). The amount of lime added to the slime (or slurry) ranges from 0.75 to 1.5 kg/t to maintain an alkalinity between 0.010 and 0.25 per cent CaO (Adamson, 1973). Subsequently, the thickened slime is pumped to the cyanidation tanks, where approximately 0.15 kg NaCN/t (or KCN) is added to dissolve the gold (Funke, 1990). In addition, compressed air is passed through the slurry to provide oxygen, which is required for the dissolution reaction:
Cyanide is consumed during agitation by the decomposition products of pyrite and the presence of CO2 contained in the compressed air. The total air-agitation for the maximum dissolution of gold varies from 15 to 45 hours, depending on (after Funke, 1990):

• Grain size of the free gold particles;
• Degree of pyrite encasement;
• Consistency  of slime grading.

Filters currently achieve separation of the gold-cyanide solution from slime. As a result of lime added to the cyanide solution, CaC03 precipitates due to the reaction with CO2 in the interstices of the filter cloth, leading to a gradual reduction of the filtration rate (Adamson, 1973). The precipitation of gold from the filtered cyanide solution is achieved by the reaction with zinc dust and the addition of small quantities of lead nitrate, which is not shown in the reaction below. The chemical reaction of the precipitation of gold from the cyanide solution can be expressed by the reaction:
The zinc-lead gold precipitate is subsequently removed from the solution by filter presses. The precipitation product or cake in the filters is passed to an acid vat where sulphuric acid is added to dissolve excess zinc and other soluble constituents. After dewatering, the slime is roasted by calcining at approximately 6000 C and is then smelted in electrode arc furnaces between 1200 and 13000 C for a period of about two hours. Finally, the recovered gold is transported in bars to the refinery plant.
The carbon-in-pulp process is applied to recover gold directly from the cyanide leached slime by adsorption onto granular activated carbon. The gold-loaded carbon is separated from the slimes by screening and is then eluted with hot NaCN under pressure to achieve a gold-containing solution. The gold can be recovered by either direct electro-winning cells or by zinc precipitation and subsequent smelting. Due to the smaller volumes used by this approach, financial savings are considerable (Funke, 1990).
Large quantities of sulphuric acid are required for the extraction of uranium from gold plant residues. Pyrite is recovered from some ore as a by-product to produce sulphuric acid. For this process, copper sulphate (CUS04) is essential for the successful froth flotation of pyrite by means of a collector (xanthate) and a phosphate containing frother (Aerofloat 25) according to Adamson (1973). In such cases the pyrite content in the tailings will be reduced, whereas the Cu and phosphates contents will be increased. In contrast to the recovery of gold by cyanide, complete dissolution of uranium is achieved by oxidizing agents such as MnOz (until about 1970) or ferric sulphate (since about 1970), the latter being produced by bacterial oxidation of ferrous sulphate. The MnOz would cause an increase of MnS04 in the tailings.

1 INTRODUCTION 
1.1 STATEMENT OF THE PROBLEM
1.2 RESEARCH OBJECTIVES
1.3 PREVIOUS WORK
1.4 STUDY AREA
1.4.1 Regional setting within the Vaal River barrage catchment
1.4.2 Climate
1.4.3 Location of the study sites
1.5 ACKNOWLEDGEMENTS
2 GEOLOGY, MINERALOGY AND CHEMISTRY OF THE GOLD ORE AND RELATED TAILINGS 
2.1 HISTORICAL AND GEOLOGICAL ASPECTS
2.2 MINERALOGy
2.2.1 Macroscopic description ofthe gold-bearing conglomerate
2.2.2 Mineralogical composition of the gold-bearing conglomerate
2.2.3 Mineralogical composition of gold mine tailings
2.3 CHEMISTRy
2.3.1 Chemical composition of the gold-bearing conglomerate
2.3.2 Chemical composition of gold mine tailings
2.4 GOLD RECOVERy
2.4.1 Metallurgical process
2.5 MANAGEMENT AND RECLAMATION OF TAILINGS DAMS IN SOUTH AFRICA
2.5.1 Introduction
2.5.2 Construction
2.5.3 Operation and decommissioning
2.5.3.1 Seepage and the development of a groundwater mound
2.5.3.2 Seepage control measures
2.5.4 Reclamation
2.5.5 Land use after reclamation
2.6 HYDROGEOCHEMICAL PROCESSES DURING THE WEATHERING PROCESS OF TAILINGS
2.6.1 Sulphide oxidation and acid generation processes
2.6.1.1 Primaryfactors
2.6.1.2 Secondaryfactors
2.6.1.3 Tertiaryfactors
2.6.1.4 Subsurface and downstreamfactors
2.7 OCCURRENCE OF TRACE ELEMENTS IN SOIL AND ITS TOXICITy
2.7.1 Occurrence of trace elements in soil
2.7.2 Toxicity
2.7.3 Environmental quality standards
3 METHODS OF INVESTIGATION 
3.1 SCOPE OF WORK
3.2 FIELD SURVEY AND SITE INFORMATION
3.3 SAMPLING AND LABORATORY TESTING
3.3.1 X-ray fluorescence spectrometry (XRF)
3.3.2 Soil extraction tests and inductively coupled plasma mass spectrometry
3.3.2.1 Method/or the soil extraction test
3.3.2.2 Inductively coupled plasma mass spectrometry (ICP-MS)
3.3.3 X-ray diffraction (XRD)
3.4 GEOTECHNICAL PROPERTIES
3.4.1 Estimation of hydrogeological conditions from geotechnical data
3.4.2 Soil types and properties
3.5 DATA EVALUATION
3.5.1 Correlation coefficients
3.5.2 Geochemical background values
3.5.3 Estimation of hydraulic conductivities in soils
3.5.4 Short-term impact
3.5.5 Long-term impact
4 CASESTUDIES 
4.1 CASE STUDY A
4.2 CASE STUDY B
4.3 CASE STUDY C
4.4 CASE STUDY D
4.5 CASE STUDY E
4.6 CASE STUDY F
4.7 CASE STUDY G
4.8 CASE STUDY H
4.9 CASE STUDY I
4.10 CASES STUDY J
4.11 CASE STUDY K
5 ENVIRONMENTAL IMPACT ASSESSMENT 
5.1 INTRODUCTION
5.2 CHARACTERISATION OF THE PRIMARY CONTAMINATION SOURCE
5.3 SHORT-TERM IMPACT ON THE SUBSURFACE
5.4 LONG-TERM IMPACT ON THE SUBSURFACE
6 RISK ASSESSMENT AND REHABILITATION MANAGEMENT 
6.1 INTRODUCTION
6.2 RISK ASSESSMENT
6.3.1 Treatment technologies
6.4 REMEDIATION OF GROUNDWATER CONTAMINATED BY ACID MINE DRAINAGE
6.5 LONG-TERM ENVIRONMENTAL MANAGEMENT FOR LARGE CONTAMINATED AREAS
6.6 ENVIRONMENTAL MONITORING AND AFTER-CARE MANAGEMENT.
6.7 FINANCIAL IMPLICATIONS OF REMEDIAL MEASURES
6.8 ENVIRONMENTAL MANAGEMENT MEASURES REQUIRED FOR THE INVESTIGATED SITES
7 DISCUSSION AND CONCLUSIONS 
7.1 DISCUSSION
7.2 CONCLUSIONS
8 LIST OF REFERENCES 
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