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The roasting of sulphide ores and concentrates is often the first step in the production of metals or chemicals. In many processes, the production of sulphuric acid is viewed as a by-product, while in some plants production is an important economic factor. Regardless of the purpose, a pyrite roasting plant consists of mainly three plant sections: roasting, gas cleaning and sulphuric acid. With the addition of air, the pyrite concentrates are transformed into solid oxides and gaseous sulphur dioxide at temperatures of 600 - 1000° C. After cleaning and cooling, the sulphur dioxide in the roasting gas is further processed to sulphuric acid. Two types of reactors are used depending on the application: stationary or circulating fluid bed . For over 60 years, Outotec has progressively been developing the principle of fluidised bed technology in several different reactor types for a multitude of process applications. The versatility of the fluidised bed reactor system has manifested itself in the treatment of minerals, including solid fuels, and for metallurgical processes both in the ferrous and non-ferrous fields. Process applications have included roasting, calcining, combustion and charring of coals, as well as off-gas treatment. This paper provides a summary of the pyrite roasting technology currently used along with a simple cost comparison of pyrite roasting and sulphur burning processes.

Jan 1, 2009

By J. M. Garcia

This paper highlights several management practices from the oil and gas industry to support the proposition that financial performance in the finite, non-renewable resource business relates more to upstream rather than downstream activities. Based on the analysis of nine oil and gas companies, this study supports a previous study involving fourteen mining companies that showed reserves growth is one of the main levers of value creation in mining. Interestingly, this study also finds that the oil and gas industry has been historically more profitable than mining. The reason, it is argued, is that oil and gas companies count on management practices that focus primarily on the upstream segments of the business, compared to the traditional downstream focus of mining. This paper delves into these ideas to conclude that what mining may need to improve its competitive advantage is a new organizational framework. Another conclusion is that the upstream management focus is vital not only for strategy formulation in the resource business, but also for policy formulation in economies based on the export of finite, non-renewable resources.

Jan 1, 2011

By S. Zhao, S. P. He, M. Y. Zhu, Y. Tao

"The effect of two refining slags (A: 1.0–1.8 CaO/Al2O3 ratio, 11–20 wt% SiO2; and B: 1.0–1.2 CaO/SiO2 ratio, 25–35 wt% Al2O3) on the inclusion morphology of a carbon structural steel (Q235B) was studied by controlling the balance between the steel and the refining slag at 1600ºC in a Si–Mo resistance laboratory furnace. The results showed that most of the inclusions in the molten steel balanced with slag A were formed by a lowmelting CaO–MgO–Al2O3–SiO2 system with a total oxygen content (T[O]) being fixed at 6–13 ppm. The inclusions in liquid steel balanced with the slag B were mostly formed by the CaO–MnO–Al2O3–SiO2 system, with T[O] controlled to 19–26 ppm. The overall size of the inclusions under balance conditions was higher in the steel reacted with the slag B than that reacted with slag A, with the latter showing more inclusions smaller than 3 um. In order to generate a large number of spherical CaO–MgO–Al2O3–SiO2 inclusions smaller than 3 ??m while decreasing the T[O] in the resulting steel, the composition of slag A should be used. This composition allowed improving the cleanliness of the slab, and is also advantageous for the castability of the molten steel. IntroductionThe sample used in this study was a typical carbon steel manufactured by CQ Steel. The galvanized, tinned, and plastic composite steel plates made of the original plates are used extensively for decorative panels, general dust pipes, drums, instrument cases, switchboxes, and train cars, among other uses. The target control composition is shown in Table I. The S and P contents are usually fixed during production below 0.020 and 0.025 wt%, respectively, with the aim of reducing longitudinal cracks on the slab surface."

Jan 1, 2017

By Willem Le Roux

Occupational health and safety is regulated by statute and common law. South Africa has, like many other countries, legislated extensively on occupational health and safety. Despite the body of legislation, the common law continues to play an important role as far as it concerns contractual, delictual and criminal liability. Two major Acts regulate occupational health and safety viz. the Occupational Health and Safety Act, No. 85 of 1993 (?OHASA?) and the Mine Health and Safety Act, No. 29 of 1996 (?MHSA?). The MHSA applies to mines and works as defined and related aspects. OHASA applies to other industries but does not apply to employers and workplaces to which the MHSA and certain matters covered by the Merchant Shipping Act, No. 57 of 1951 apply. See section 1(3) of OHASA. 2 The employer?s health and safety obligations in terms of the MHSA 2.1 Section 2(1) of the MHSA provides amongst others, as follows: ?The employer of every mine that is being worked must: (a) ensure, as far as reasonably practicable, that the mine is designed, constructed and equipped: (i) to provide conditions for safe operation and a healthy working environment (ii) ... (b) ensure, as far as reasonably practicable, that the mine is commissioned, operated, maintained and decommissioned in such away that employees can perform their work without endangering the health and safety of themselves or of any other person?.

Jan 1, 2011

By P. J. How, T. O. Hagan, M. F. Handley

Inelastic deformation of highly stressed rock in deep mines needs to be quantified in both time and space. Since most deformation occurs by shear, continuous monitoring of discontinuities in the rock from boreholes holds promise. The casing. Al interpretations of shear movements are based on a simple relation between shear displacement and the resulting change in tilt of the sensor. A prototype has been constructed and has undergone laboratory trials with satisfactory results. Underground trials will commence shortly. The instrument as designed has the potential for other applications, both in the civil-engineering and mining industries. paper describes an instrument that is both cheap and reliable, and capable of monitoring such shear movements continuously. The principle of operation is based on two capacitive gravity sensors mounted orthogonaly to each other in a sealed aluminum

Jan 1, 1988

By Fadeela Saloojee, Frank K. Crundwell

"Pressure leaching is an option for copper recovery from chalcopyrite. Leaching takes place at high temperatures and pressures in the presence of an oxidising agent. This work shows how the performance of pressure leaching circuits may be improved by optimising the configuration of autoclaves and the heat-removal method. The objective was achieved by creating an autoclave model by combining mass balances, energy balances and population balances. A base-case circuit consisting of a single autoclave was chosen, and this base case was compared with options for increasing capacity by adding more autoclaves to the circuit. These capacity–increase options include circuits in which the additional autoclave is added in parallel, in series and in series with thickening between the two autoclaves. The copper extraction, productivity and cooling requirements for these options are compared. The series circuit with inter-stage thickening has the highest extraction and productivity, however, the cooling duty in the first autoclave is high. In addition to comparing circuit options, three options for heat removal were investigated: cooling coils, quench water and flash recycling. The flash-recycle option results in the highest copper extraction, and the quench-water option gives the lowest copper extraction. INTRODUCTION Background Copper sulfide ores and concentrates containing chalcopyrite are typically hard to leach under atmospheric conditions. Smelting of these ores may also be a challenge, due to the inability of smelting processes to handle low-grade ores, as well as impurities such as arsenic and antimony. Alternative hydrometallurgical routes to treating chalcopyrite are therefore required. One option is pressure leaching, which is the focus of this work. In the pressure leaching process, sulfide ores and concentrates are leached in autoclaves at temperatures of 200 to 220°C and a pressure of about 3000 kPa (Schlesinger et al., 2011). An oxidising agent, usually oxygen, is required to oxidise sulfide ions to sulfate ions. The oxidation of sulfur to sulfate is exothermic; consequently, a method of heat removal is necessary in order to maintain the temperature in the autoclave."

Jan 1, 2015

By S. Lubbe

South Africa and Australia are the biggest suppliers of zircon sand to the international zirconium industry. However neither South Africa nor Australia is well known for zircon beneficiation. Geratech Zirconium Beneficiation Ltd (GZB) continued with additional research on sodium hydroxide (NaOH) cracking of zircon sand during 2002-2003. In 2003 GZB started extracting zirconium from zircon sand by means of NaOH cracking on a commercial scale. Experience has shown that temperature profile and atmospheric control inside the furnace is crucial for the beneficiation of zircon sand. Silica carryover to zirconium chemicals could result if a high temperature is used. Once the sodium silicate is extracted from the sodium zirconate and dissolved in hydrochloric acid, two distinct routes can be followed to precipitate various zirconium chemicals. The most common route is to precipitate zirconium oxychloride crystals (ZOC), with subsequent purification from all contaminants (crystal route). Less known is the process (liquid route) that involves the direct precipitation of zirconium basic sulphate (ZBS). This route will yield a less pure product, with contaminants such as silica and titanium. An important factor in this route is the prevention of silica gel formation, which could hamper final product filtration. For applications like paint drying (zirconium octoate) or antiperspirants (zirconium hydroxychloride) low levels of contaminants have no effect on the final product. The advantage of the liquid route is fewer production stages compared to the crystal route. The disadvantage of the liquid route is that the market for the products will be significantly smaller. The optimum solution could be a plant design that could cater for both routes. Another example of an application of zirconium chemicals is the use of ammonium zirconium carbonate (AZC) in the paper industry. Zirconium basic carbonate (ZBC) is dissolved in ammonium carbonate to produce AZC solution. AZC is used mainly in European countries in the paper industry. For example, carton boxes were initially produced with formaldehyde as the binder, however, it has now been replaced with AZC since formaldehyde is considered toxic. AZC reacts with the cellulose fibres in the paper to act as the binder. The resulting product is not toxic, and printing ink dries very quickly due to the porous paper structure. Other applications of zirconium chemicals involve the use of acid zirconium sulphate tetrahydrate (AZST), zirconium orthosulphate (ZOS), potassium zirconium carbonate (KZC), and zirconium hydrous oxide (ZHO). Fluoride-based zirconium chemicals like zirconium tetra-fluoride (ZrF4) and hexafluoro-zirconic acid (H2ZrF6) are used in the aluminium industry. Zirconium oxide (Zr02) can be produced from any of the abovementioned precipitated chemicals via a high-temperature decomposition process. The physical properties of such oxides can differ tremendously, depending on the final application. The fired density of sanitaryware is typically 5.8 g/cm3, compared to milling media at >6.1 g/cm3. The required fired density is achieved by controlling the precipitation and decomposition conditions for these two oxides. The same applies to stabilized and mixed oxides, for example where zirconium oxide acts as an oxygen carrier in fuel cells.

Jan 1, 2012

By T. Käyhköa, J. Liipoa, S. Khizanishvilic, M. Sinche-Gonzalezb

The use of geometallurgical modelling is becoming more common in concentrator plant design and during the operational phase. Predictive modelling aims to define optimised process parameters for the concentrator plant to accommodate variability within the ore feed. The main objective of this study is to show how the metallurgical response of different types of ore blends can be simulated using kinetic models derived from experimental flotation tests with different geometallurgical end-member ore types and their blends. For these validation tests, four different end-member ore types containing different ratios of chalcopyrite, chalcocite and oxidised copper minerals from Rich Metal Group’s Madneuli copper-gold mine, Georgia, were tested. Kinetic flotation tests, including both rougher and cleaner flotation tests, were carried out. The simulations were performed using HSC Chemistry software based on mineral compositions of the ores and kinetic models using rectangular distribution equations and flowsheet model. Simulated and experimental results of copper grade, copper recovery and concentrate mass pull correlated well with R2 values of >0.95. Based on these results, the metallurgical response of different ore blends can be simulated based on the flotation kinetics of end-member ore types. The use of this kind of simulation will create savings during the test work phase and increase the flexibility and predictability in the processing stage. Keywords: Ore body characterisation, geometallurgy, process mineralogy

Jan 1, 2020

By D. A. J. Ross-Watt

Right now the South African mining industry is a high-profile, vital contributor to the nation just as it has been for the past one hundred and twenty-six years since the discovery of diamonds-perhaps even further back if the mining of copper, iron ore, tin, and other minerals in earlier centuries is included. The media in 1995, as has been the case over the years, bear testimony to the complex interactions between the industry and its stakeholders: shareholders, government, employees and their unions, training and educational institutions, researchers, markets, the public, and suppliers of goods, services, and infrastructure. The pressures and challenges that the industry continues to face can be summarized as safety and health, competitiveness, and social and environmental responsibility. There are no surprises here. The motivation for the industry to perform well in these areas is clear; the focus I intend to bring is on the most important tool required for success-that of good mining engineering. This Address sets out - to review the mining aspects that have been discussed in South Africa during the first five years of the decade, a discussion that shows the return of the country as an important mining forum, confirms the role of professional institutes and associations in the development of mining engineering, and brings current relevant issues into focus - to highlight the vital role of mining engineering in the industry, and to identify priorities for the remainder of the decade - to provide the prospective mining engineer, the mining-engineering student and trainee, and the practising mining engineer and mine manager with a better appreciation of the mining-engineering discipline. institutes and ssociations

Jan 1, 1995

By H. P. Kritzinger, T. C. Kingsley

Rotary kilns are used for a variety of mineral processing operations. Hatch makes use of a kiln model developed from first principles to evaluate designs for its clients. This tool has been applied to a variety of pyrometallurgical applications. This paper illustrates the application of numerical optimization techniques in combination with the kiln model in the interrogation of a generic iron-ore reduction process. The fundamental modelling concepts are explained, followed by a description of the optimization approach. The results show how the combination of the two methods, with modern computing power, can generate a large number of viable design and operating candidates.

Jan 1, 2014

By D. Skelton

The Renard kimberlite cluster is located in the Monts Otish region of Quebec, Canada. A Canadian National Instrument (NI) 43-101 compliant resource statement for the Renard kimberlites 2, 3, 4 and 9 and the Lynx kimberlite dyke was first issued in 2008 followed by a Preliminary Economic Assessment and the development of a conceptual mine plan. Following a successful drill program in 2009 that greatly expanded the amount of kimberlite in Renard 2, a revised resource statement was issued in December 2009 comprising 23 million carats of Indicated Mineral Resources and 13 million carats of Inferred Mineral Resources, a threefold increase on the previous estimate. The project is currently the focus of a second Preliminary Economic Assessment and it is expected that a full feasibility and permitting program will commence in 2010 leading to the creation of Quebec's first diamond mine by 2013.

Jan 1, 2010

This paper presents an initial investigation into the viability of co-kriging the grades of multiple mining cuts in a two dimensional, thin reef scenario. Deposits with gradational mineralized zones, in the perpendicular sense, are not as well served by the use of reef accumulation and thickness as deposits characterized by a sharp distinction between ore and waste rock. South African Bushveld Complex platinum deposits and Witwatersrand gold deposits are examples of this distinction. The calculation of an optimal mining cut is of paramount importance in the mining of Bushveld platinum. Two-dimensional kriging of a constant stoping width mining cut is an effective method for resource calculations once a cut is decided, but this begs the question of how to decide on the cut. Comparison of multiple cuts is used as a way of achieving this decision on many of Anglo Platinum’s operations. Although the kriging of each cut is valid in isolation, occasional negative values occur when incremental grades between overlapping cuts are calculated. Negative incremental grades cannot reflect reality. They are thought to be related to the inability of the method to include the strong continuity in the perpendicular direction demonstrated by the relevant variograms. Co-kriging of related variables has been investigated as a potential means tore impose this inherent order relationship. Various co-kriging options have been investigated. The method has been tested on a real data set from an established platinum operation. One of three types of anomaly has been investigated so far. The results show that co-kriging the mining cut grade as the primary variable, with just one secondary variable, the reef thickness, eliminates most occurrences. Severe anomalies remaining are arguably indicative of population mixing and might be more effectively treated by rezoning the data. Aim—To investigate the viability of co-kriging the grades of multiple mining cuts as a method for resolving anomalies in the estimation of such cuts on Bushveld platinum deposits, with applicability to other thin reef type deposits with a gradational not sharp cut-off between ore and waste in the direction perpendicular to the reef. Keywords: co-kriging, cross variograms, ordinary kriging, platinum, two-dimensional, grade profile

Jan 1, 2003

By Y. G. Zhu, G. B. Zheng, X. X. Hu

Reverse flotation by cationic collectors is the classical method to reduce the silicon content in magnesite. But because of the low selectivity, there is a large loss of magnesite in the reverse flotation froth. To improve the efficiency of the collector, a novel collector named BK433 was developed in this paper. In order to determine the performance of the BK433 collector, a systematic experimental study was carried out with a high silicon magnesite in China. The suitable pH environment of slurry was confirmed, the depression effect of different depressants was investigated, and the flotation effect was compared with commonly used traditional collectors. The results show that the BK433 shows remarkable good selectivity and collection ability. After the 4 stages reverse flotation, the SiO2 content decreased from 3.76% to 0.25%, SiO2 impurity removal rate reached 94.69%. Meanwhile, the MgO content increased from 45.06% to 46.73%，the recovery of obtained concentrate (calculated by MgO) was 82.79%. Keywords: Magnesite, Desilicication, Reverse flotation, Collector

Jan 1, 2020

By D. K. Chibwe, P. Taskinen

"Typical operation of industrial Peirce-Smith converters (PSCs) used in copper smelting results in several common phenomena such as excessive splashing, slopping, tuyere blockage, and low converter campaign life due to high tuyere line refractory erosion. These phenomena give rise to an inefficient operation from both a process and an energy perspective.This investigation presents numerical and physical models of flow patterns, mixing, solid-liquid mass transfer, and slag-matte phase distribution in a slice model of an industrial PSC. For physical simulations, a 0.2-scale cold model was developed. Water, kerosene, air, and sintered benzoic acid compacts were used to simulate matte, slag, injected gas, and solid additions into the PSC. The two- and three-dimensional numerical simulations of the three-phase system were carried out using volume of fluid (VOF) and realizable k- e (RKE) turbulence models to account for the multiphase and turbulence nature of the flow respectively. These models were implemented using the commercial computational fluid dynamics numerical code FLUENT.Both numerical and physical simulations were able to predict, in agreement, the mixing and dispersion characteristics of the system in relation to various blowing conditions employed in the investigation. Measurement of mass transfer characteristics conclusively indicated that fluid flow in the PSC is stratified. Both blowing configuration and volume of slag in the converters were found to have significant effects on mixing propagation, wave formation, and splashing.The splashing and wave motion in these converters are known to cause losses of metal/matte and potential production time due to the requirement for intermittent cleaning of the converter mouth, resulting in reduced process throughput. To prevent these losses and hence to increase process efficiency, we propose a combined blowing configuration using top lance and lateral nozzles. The numerical simulations were conducted on combined, as well as lateral, blowing conditions. In comparison, the recent results of the combined blowing concept were found to be encouraging."

Jan 1, 2014

By T. Szendrei, S. Tose

The generally accepted view in rock blasting is that the sources of energy for the fracture and movement of rock reside in the shock wave and gas action resulting from the explosion, and yet the mechanisms by which these sources interact with the rock have remained unclarified. It has also been noted that up to 50% of the work capacity of an explosive released in a blast cannot be accounted for by field measurements of energy partitioning. In this study, we describe a physical model that details the response of rock to both shock wave and gas action. An analytical model based on momentum conservation is derived to describe the dynamics of shock-driven expansion of the blasthole. Radial expansion of the hole is the key parameter that permits the derivation of the following characteristics of rock response to shock loads: hole expansion time; volume of displaced rock; energy consumed per unit volume; expansion energy efficiency; stress wave pulse length; gas pressure in enlarged hole. Soon after the completion of hole expansion, the shock wave degenerates to an elastic stress wave that runs through the burden. Blasthole expansions of between 50% and 300% of diameter are completed in under 1 ms and, depending on rock properties, consume 32% to 42% of the detonation energy or about 55% of the available mechanical (Gurney) energy. Gas pressure in the enlarged holes in five rock types is between 35 MPa and 650 MPa, and drives the mass movement of burden rock.

Oct 9, 2024

By Willem Le Roux

1.1 Section 47(1) of the Mine Health and Safety Act, 29 of 1996 (?the MHSA?) establishes a Mine Health and Safety Inspectorate which is administered by the Chief Inspector of Mines. The Mine Health and Safety Inspectorate is part of the Department of Mineral Resources. The Chief Inspector is appointed by the Minister of Mineral Resources (see section 48(1) of the MHSA). 1.2 The Chief Inspector must, amongst other things, ensure that the provisions of the MHSA are complied with and enforced, and that all the duties imposed upon the Chief Inspector and the other Inspectors of Mines in terms of the MHSA and any other law are performed. The Chief Inspector of Mines must appoint Principal Inspectors of Mines and Inspectors of Mines to assist him/her to discharge his/her duties (see section 49(1)(a) and (c) of the MHSA).

Jan 1, 2011

By R. Abela

In over ninety per cent of mineral sand plants, the most commonly used equipment is the spiral concentrator. The tonnage treated over typical spiral plants can range from as low as 200 ton per hour up to 6 000 tons per hour. Due to the limited capacity of each spiral, plants consist of multiple spirals assembled into banks. The mixture of minerals present in the sand feed requires several spiral stages to achieve the desired mineral recovery. Due to the complexity of the spiral circuits, there are always some design and operating problems that present themselves in plants. These range from slurry distribution, spirals operating at incorrect design parameters and feed variations. The material being treated is difficult to maintain in suspension due to its high density, resulting in segregation and preferential flow in pipelines, distributors and sumps. The objective of this paper is to highlight common problems associated with spiral installations to designers and operators of plants. Practical solutions are recommended in order to prevent and minimize these mistakes in the design. These solutions may be used to improve performance on operating plants.

Jan 1, 2003

By J. O. Claassen

"SynopsisMining value chains are dynamic systems and should be managed according to systemic flow-based principles. This includes a detailed focus on the impact of variable geological conditions on material flow, product quality, and production cost, visibility of the total mining system to operators and managers, as well as the synchronization of resources and activities in downstream processes. Thirty mining operations across the African continent were studied to establish to what extent systemic flowbased principles are applied in day-to-day operations. The study indicated that although most mining operations have identified the relevant geoprocessing variables, only a small number apply this information in a flow-based management approach. Most mining operators do not focus on developing a clear flow view of the mining value chain and on making it visible to operators and managers to enable them to optimally set and reset operations. Daily synchronization of geoprocessing variables is limited to a small number of variables, e.g. synchronizing ore hardness with drilling or milling requirements, in less than 30% of the operations evaluated. Operations where these principles have been adopted reported a noteworthy improvement in performance. The study indicates that notable potential exists in most of the operations evaluated to implement flowbased management principles, including geometallurgical principles. This can significantly improve value chain performance without exorbitant capital layouts and enhance ROCE. IntroductionObjective of the studyThe study seeks to establish to what extent systemic flow-based principles have been adopted in the southern African mining industry and the implications thereof. Specific attention is given to the importance of defining and managing geoprocessing dependencies (geological variables directly impacting mining and plant performance – the geometallurgical approach), establishing a flow view of the mining system, which typically includes a dashboard of the flow of ore/product quality and cost from the mining face to the customer, and limiting the impact of dependencies through synchronization of activities and resources, e.g. ensuring that a continuous miner does not wait for a roofbolter, or that reagent addition rates are matched with a specific ore/concentrate composition."

Jan 1, 2015

By C. Banff, E. C. F. Hamman, J. Venter

Open pit slope stability analyses for large scale instabilities are often carried out using limit equilibrium, finite element or finite difference computer programs. All of these require empirical estimates of rock mass shear strength when calibrated values are not available, which is a common occurrence. A common method of estimating rock mass shear strength is the Hoek-Brown failure criterion, which often over- or underestimates rock mass shear strength when used in limit equilibrium or elastic linearly plastic slope stability analyses, resulting in published alternative estimates for the coefficients of the Hoek-Brown failure criterion. This paper investigates the differences between limit equilibrium and elastic linearly plastic slope stability analysis based on the Hoek-Brown failure criterion. The investigation consisted of a comparison of 3840 RS2 finite element analysis models completed with strain softening, and 216 Slide2 limit equilibrium models. The models were used to evaluate a methodology to determine equivalent limit equilibrium and elastic linearly plastic shear strength parameters that provide the same stable slope angles and instability back break distances as the more representative strain softening stress analysis methods. This paper presents the methodology and compiles the results into design charts and equations for equivalent limit equilibrium Mohr-Coulomb shear strength parameters, that simplify calibration against actual case studies. The results compare well with older empirical shear strength estimates, however, a detailed calibration of the results against case studies will need to be carried out. The analysis also did not consider pore pressure effects; this is left for future publication.

Sep 1, 2025

By B. R. Broekman, D. W. Penman

Discussion by P.J Lloydt Broekman and Penman describe (J.S. Afr. Inst. Min. Metal., vol. 91, Aug. 1991, p. 261) the addition of sodium cyanide and zinc sulphate to the flotation feed at Prieska. They note that for 'safety sake, lime was added to the rod mil to raise the pH to 9,5', and that 'personnel working in the vicinity of the rod mil were required to wear gas masks'. Author reply to P.J. Lloyd We thank Dr Lloyd for his interesting contribution. As he correctly points out, the Prieska float had run with cyanide solutions at pH values of 6,4. In fact, it ran safely for extended period with the pH slightly below 6.

Jan 1, 1991