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By G. M. Evans, S. Mitra, G. J. Jameson, M. M. Hoque, E. Doroodchi, R. Moreno-Atanasio

In this study, the detachment behaviour of 314 μm diameter silica particles from a single 3 mm diameter bubble rotating in a confined cavity connected to a cross-flowing water channel was investigated. The liquid cross-flow generated a rotating flow inside the confined cavity, which was quantified using 2D particle image velocimetry (PIV) measurements. The fluctuating (x-y) velocity components inside the cavity for each cross-flow velocity, U, were then used to estimate the corresponding time/volume average turbulence intensity values, Ti. Turbulence intensity ranged from 9.3-16.2%. In the absence of a bubble, the rotating vortex within the cavity created a pressure field which decreased towards the centre of the vortex; wherein the lowest pressure zone was located at the centre of the cavity. The vorticity into the cavity was found to increase with increasing the turbulence intensity. A bubble, both without and with particles attached, was introduced into the cavity under different flow conditions and its motion was tracked using the shadowgraphy technique. Experimentally, it was found that the bubble-particle aggregate was trapped at the centre of the cavity (low-pressure region) following a spiral trajectory. An interface resolved 3D computational fluid dynamics (CFD) model was developed to predict the particle detachment conditions within the confined cavity. CFDpredicted vorticity in the absence of a bubble agreed well with experimental observation. The CFD analysis indicated that the vorticity inside the cavity was not significantly changed when a single bubble was introduced. On this basis, the computed vorticity was utilised in describing the force balance applied to the attached particle. Computationally, it was found that the critical turbulence intensity, beyond which the particle was detached from the bubble, was 14%. This value was consistent with the experimental observation of detachment at Ti at 15.7%, and above, but not at lower values. Keywords: Bubble, particle, turbulence intensity, vorticity, particle detachment, CFD, PIV

Jan 1, 2020

By M. A. Belokrys, F. O. Milovich, D. D. Aksenova, V. A. Ignatkina, V. A. Bocharov, A. A. Kayumov

Copper remains one of the most sought-after metals in the industry. Tennantite (Cu12As4S13) and tetrahedrite (Cu12Sb4S13) that occur in pyrite copper, copper-zinc and polymetallic ores are an additional reserve for increasing technological parameters of flotation. Pyrite mass (85% and more) has similar flotation properties to other sulfide minerals, and imperfection of reagent regime is a common problem that occurs during quality copper concentrate processing. Due to tennantite's and chalcopyrite's differences in flotation properties with butyl xanthate at pH > 10, tennantite often reports to the flotation tails. Hight alkalinity of the medium increases the oxidation level of sulfide minerals, thus increasing the concentration of S2O3 2-, SO3 2- SO4 2-. At processing facilities in the Russian Federation, the concentration of thiosulfate ions in pulp is in the range of 0.5 to 1.5 g/l and higher, which leads to a decrease in flotation activity of the tennantite. The effect of butyl xanthate (as standard), thionocarbamate, dithiophosphate and M-TF (a mixture of dithiophosphate and thionocarbamate) on monomineral (tennantite and pyrite) flotation in the presence of sodium thiosulfate was studied. The variable parameters of the "thiosulfate-collector" flotation complex were thiosulfate concentration (100-900); pH (8-12,5). The scientific methods included contact angle for hydration assessment; TEM for surface studies; UV spectroscopy and iodometric titration for solubility control. Comprehensive studies have established a connection between the concentration of thiosulfate ions and tennantite flotation selectivity. Flotation activity of tennantite is higher at pH 8 with a thiosulfate concentration under 300 mg/l. Pyrite remains active with butyl xanthate under the same conditions. TEM images of the surface films of the collectors on pyrite were acquired at pH 8. The MTF provides the biggest difference in adsorption, wettability and flotation activity between tennantite and pyrite at pH 8 and thiosulfate concentration under 300 mg/l, in comparison with butyl xanthate. The results of the research will be applied to improve selective reagent regimes of tennantite minerals flotation. Keywords: Tennantite, pyrite, chalcopyrite, sphalerite, flotation, sulfhydryl collectors, thiosulfate, adsorption, mechanism, reagent regime

Jan 1, 2020

By G. W. Holl, E. G. Fairon

A review of some aspects of shaft design by G. W. HOLL, Pr. Eng., Associate of the Mine Managers Association of South Africa (Fellow)- and E. G. FAIRON, Pr. Eng., F.S.A.I.E.E., M.I. Cert. M.E.E., M.I.Mech. E., (Visitor) The reasons for the choice of vertical shafts in preference to inclined shafts of comparable capacity are set out, and the conclusion that the former are superior in virtually all circumstances is substantiated. The shapes of vertical shafts sunk during the past sixty years are recorded, and the shapes of those sunk during the past decade are analysed. From this analysis it is clear that circular and elliptical (and quasi-elliptical) shafts are the shafts of the future and that the rectangular shape will conceivably become obsolete. The determination of the size of shaft is discussed and it is forecast that shafts of a diameter of 13 metres and over will probably be sunk without difficulty in the years to come. The depth of the shaft is determined intrinsically by the Mines and Works Regulations applicable to factors pf safety of winding ropes and the design of the winding system. The Blair winder is the most suitable winder for hoisting rock, men, and material from a depth of, say, 3000 metres on an economic basis. Tertiary shafts will, therefore, be unnecessary in future deep-level mining projects at the limit of depth now envisaged. A few other factors relating to shaft design are described.

By P. van der Heever

The genesis of Simrac?s technology transfer practices and future initiatives are discussed in this report. Simrac produces a relatively large volume of research-based health and safety outputs that must be communicated to its tripartite stakeholders and to other interested and affected parties. The wide spectrum and complexity of the transferables, together with the diversity of end-user needs and skills of the recipient population, necessitate customizing of technology transfer practices. In an attempt to optimize this process, a variety of practices are employed. Information is transferred by means of publication and dissemination of a variety of reports (paper, website and CD).Knowledge transfer is directed to health and safety practitioners by means of workshops, seminars, best practice booklets, handbooks and textbooks. Promotion of recently completed projects is directed at targeted recipients by means of project launches. Only isolated cases of piloting and implementation of technologies have been undertaken due to lack of resources and past organizational focus. It is anticipated that Simrac?s new organizational structure and management system, together with end-user involvement by industry champions, from project inception to implementation, will enhance the effectiveness of the technology transfer process.

Jan 1, 2002

By N. B. Strydom, P. L. Jooste, H. F. Kotze, W. H. Van Der Walt

Note: A revised method of heat acclimatization by N. B. STRYDOM, W. H. VAN DER WALT, P. L. JOOSTE, and H. F. KOTZE INTRODUCTION Climatic-room acclimatization was introduced in the gold-mining industry in 1965 and was rapidly adopted by all mines because it reduced the number of days required for heat acclimatization from 12 to 8. The method has been adequately described in previous papers1, 2 but, briefly, requires four hours of heat exposure per day while the work rate is progessively increased from 35 W on Day 1 to 70 W on Days 7 and 8. Several attempts to shorten

Jan 6, 1976

By A. P. Squelch

The question of what stress condition exists in the hangingwall of tabular stopes remains topical and largely unanswered. Limited field data exist to quantify the situation and provide calibration for investigative computer modelling exercises. Work conducted in1989 and 1990 under a Chamber of Mines Research Organization (now the CSIR: Division of Mining Technology) programme of research into the behaviour and benefits of backfill may assist with this situation. The original work was motivated by the premise that the improved hangingwall stability in backfilled panels (identified by rockfall accident analyses) is due to a change in the hangingwall stress regime. Current, largely theoretical, studies into mechanisms of stope hangingwall behaviour and support interaction could benefit from considering these data. Keywords: Backfill, horizontal stress, stope hangingwall, hangingwall stability, support interaction.

Jan 1, 2002

1. Book reviews Stainless Steels '84. London, Institute of Metals, 1985. £39 (U.K.), US$62 (elsewhere) Reviewer: J.B. Oalton Process engineering of size reduction: bal miling, by LG. Austin, R.R. Klimpel, andP.T. Luckie. New York, AIME, 1984. 561 pp. Reviewer: M.J. Lazer 2. Recent publications Hard and superhard materials. New York, Worltech Report Inc. (44 Woodbine Avenue, P.O. Box 212, Northport, 11768 U.S.A.), 1985. 270 pp. $1500.International comparison of standard materials-steel and cast iron, compiled by W. Oto and K. Schaning, and edited by DIN German Standards Institute e.V. Berlin, Beuth Verlag GmbH, 1985. 3rd edition, 416 pp. DM 158. Tin and its uses no. 146. Greenford (England), International Tin Research Institute, 1985. Write your own CV, by Max Kowen. Johannesburg, Legal Personnel Selections (404 Medical Centre, Cor Jeppe and Delvers Sts, Johannesburg 2001), Nov. 1985. 40 pp. R12,50 + GST.

Jan 1, 1986

By W. A. Hustrulid

CONTENTS PART I The theory of percussive rockdrill operation. Blow frequency Piston impact velocity and blow energy Thrust requirements Prediction of penetration rate Drillsteel life Bit wear PART II Measurements of percussive rockdrill performance Thrust-penetration rate measurements Blow frequency Piston impact velocity and blow energy Prediction of penetration rates The effects of hole diameter Thrusts applied by men Bit wear Appendix I Experimental procedure and rockdrill and drill steel specifications Appendix II Piston impact velocities and energies PART III Practical implications. Provisions of adequate thrust Conclusions INTRODUCTION In South African gold mines some 100 million tons of rock are broken each year by drilling and blasting. Only minor improvements to this method of rockbreaking have been made since the general introduction of pneumatic rockdrills for stoping 50 years ago and, more recently, of drillsteels tipped with tungsten carbide. The combined total cost of drilling and blasting amounts to almost R0,50 per ton of rock broken, which is just short of one quarter of the direct cost of stoping or development. Rockdrilling accounts for about R0,25 of the total cost of rockbreaking, of which RO,12 represents labour costs and R0,13 the cost of drillsteels, rockdrill spares and power. Mine air compressors use over 7 X 108 kWh of electric energy a year, amounting to 7 percent of the total electric power used by the gold mining industry. The efficiency of percussive rockdrilling is, therefore, of considerable significance to mining costs. Over the past four years the Mining Research Laboratory of the Chamber of Mines of S.A. has been engaged in experimental studies of percussive drilling. These experiments were conducted at test sites on Robinson Deep Gold Mine and Vlakfontein Gold Mine, made available by Gold Fields of South Africa, Limited. Three rockdrills, designated types A, Band C, of different manufacture, were used in the experiments which involved the measurement of the following operating characteristics: 1. Throttle air pressure 2. Applied thrust 3. Penetration rate 4. Air consumption 5. Blows per minute 6. Drill steel revolutions per minute. The results of these investigations are presented in three parts. Part I, "The theory of percussive rockdrill operation" which provides a concise, mathematical description of the performance of a percussive rockdrill in terms of its principal mechanical dimensions and operating conditions, and an analysis of the bit and gauge wear of drillsteels. Part II, "Measurements of percussive rockdrill performance" in which the experimental results are presented in terms of thrust-penetration rate curves at different throttle air pressures and throttle pressure-air consumption curves, which are shown to agree well with the predictions of theory. In addition, data relating gauge wear on the drillsteel and penetration rate to the distance drilled are discussed. Part III, "Practical implications" in which an analysis of the status of thrusting devices is made and the benefits that would derive from using uptimum thrusts are discussed.

By D. A. Singer, J. C. Griffiths

"One of the most useful approaches to a complex problem is to model the system which gives rise to the problem and to perform simulations through the model to find appropriate solutions. The Engel Model may be used as such a simulator to investigate the problem of the exploration for natural resources.The Engel Model is based on a two-stage search procedure in which a known area (A) which contains m prizes is subjected to search by n coverings. During the first-stage search the sensor system records signals which may be a response from prizes (that is, true signals) or may be false. Signals tend to cluster around a prize and a cluster size consists of s signals. If a circle of radius R is drawn around each signal some will overlap and the overlap from s signals contains a prize. The cost of a single covering in which each point of the area comes under observation once, is designated E1. Second-stage detailed search is expensive and is confined to areas of overlap of s circles; it leads to the location of prizes. Second-stage search of the whole area costs Ea units. The expected value, V, of a prize is known and its development cost is E3 units.The sensor system characteristics are defined by the circle of radius R around a signal, which should be small compared with the total area A and enters the model as a parameter a x R2/A. The probability of a false signal during a single covering of a circle of radius I is a known constant, A.This model was applied to the search for uranium ore bodies using two different types of occurrence. In the first, the Wyoming Basin type, the area is relatively large (A = 1 496 square miles) and a few (m = 4) large prizes occur (expected value V = $30 million). In the second area, which represents the Ambrosia Lake region of New Mexico, the area is smaller (A = 212 square miles) and contains many (m = 200) smaller prizes (expectedvalue of a prize V,$2 million). The three costs, E1, E2, E3, are adjusted to the different circumstances in the two areas.Simulations were performed using n = 1 to 40 coverings with cluster sizes s = 0 to 19; the radius R was varied from 0.52 to 3 miles (and, therefore, a = 0.002 to 0.133 0, respectively, dependent on the area of interest). The value of A was varied from 0.01 to 0.10.Profitable programs were found over a wide range of the parameters; the two areas required different strategies to achieve maximum profit. In the Ambrosia Lake region a few passes (n = 4 to 6) with small cluster sizes (s = 2 to 6) are more profitable, whereas in the Wyoming Basin region a large number of passes (n = 35 to 40) and cluster sizes of s = 5 to 10 are most profitable.As pointed out by Allais (1957) the larger prizes are the most important in deciding the successful outcomes of exploration programs, and the profit margin in the search for a few large prizes is sufficient to yield success over a wide range of programs. Such programs require very large investments if they are to be successful.To refine the Engel Simulator as a decision-maker in selecting the most profitable exploration programs it is necessary to obtain firm estimates of the sensor system characteristics, that is, of R, a and ? (lambda)."

Jan 1, 2014

By K. Shame, H. H. Thiele

South Africa holds approximately 70% of the world’s total chrome reserves. Chromium is an important raw material for the production of stainless steel and ferrochrome alloys. Over the years, China’s ferrochrome market has been dependent on chrome ore imports from South Africa. In 2018, China imported about 75% of its chrome ore from South Africa. With the low platinum price and the rising demand for stainless steel, a growing demand to evaluate new beneficiation technologies for fine chromite was identified. Gravity separation of fine chromite (<75 μm) is inefficient, due to the limitations of the technology to upgrade fine particle size distributions. Potential for the recovery of chromite, specifically fines (<75μm), lost to UG2 PGM flotation tails, but also from MG and LG6 spiral circuit tailings exist. Flotation of podiform chromite ores from India and Kazhkhstan is successful in generating saleable product grade, following acid or alkaline flotation routes. Application to South African stratiform chromite is being evaluated, with the emphasis on evaluating reagent type, optimising dosages and the effect of pH in optimising flotation performance. Flotation can unlock additional value from chromite tailings material. The effect of ore characteristics such as gangue mineralogy, particle size distribution and headgrade are reported. The performance in terms of product quality and yield are evaluated and compared. The impact of this paper will not only prolong life of mine and increase revenue but will also alleviate environmental burdens associated with tailings dams. Keywords: Chromite, flotation, fine material, performance optimisation

Jan 1, 2020

By R. S. Young

There can be very few members of the Institute who failed to appreciate this extremely interesting and informative paper by Mr Rohrmann of Highveld Steel and Vanadium Corporation. This story of vanadium in South Africa has a very wide appeal to members of our profession: it embraces both ferrous and non-ferrous metallurgy, and both hydrometallurgy and pyrometallurgy. In a comparatively short time, South Africa has become one of the the world&apos;s leading producers of vanadium.

Jan 1, 1985

By H. Immink

Keywords: Kyoto Protocol, carbon credits, global warming, energy, auditing Climate change is nowadays being taken increasingly seriously, and the link between global warming and the carbon dioxide released in the generation of energy is coming under increasing scrutiny. The Kyoto Protocol was signed in 1997, after five years of negotiation. So far, about 84 countries have ratified this agreement, under which terms developed countries have strict targets linked to their 1990 level of emissions. However, ?greenhouse gas? (such as carbon dioxide) emissions continue to rise alarmingly. According to the World Resources Institute, greenhouse gas emissions in the USA increased 13% from 1990-2002, while Australia recorded a 22% increase. In contrast, the European Union, a Kyoto signatory, has seen a small decline in its emissions during the same period. This is led by the UK and Germany, who reduced their emissions by 15% and 19% respectively. The concept of ?carbon credits? provides a funding mechanism for companies in the developed world to compensate for some of their CO2 emissions by paying companies (trading credits) elsewhere in the world to undertake projects to bring about reductions in emissions. In the southern African mining industry, it will be possible to get refunded for undertaking voluntary emission-reduction projects. Concern about the environment is creating business opportunities. A preliminary calculation has shown that various projects in South Africa (those in the pipeline, with quite a high likelihood of coming off) could earn a combined total of about R2.5-billion, under the initial phase of the Kyoto Protocol&apos;s carbon-credit scheme, which allows developed-country companies to trade credits with developing countries. A number of local companies are planning projects that could comply with the Clean Development Mechanism (CDM), in which they will implement techniques to achieve certified emission reductions (CERs) of greenhouse gases. In order to prevent abuse of such a system involving vast sums of money, vigorous audits need to be carried out. A company has to prove it has made real changes and is not simply conducting business as usual under a cloak of claimed environmental improvements. PricewaterhouseCoopers has applied to become one of only a few international agencies that will be accredited to conduct validation and verification of the credits, by auditing what is called the ?designated operational entity?. An independent validator needs to confirm that the project meets the strict requirements of the Kyoto Protocol. The company has to demonstrate that it?s not business as usual. So, if you refurbished every five years, the next refurbishment is not a good CDM project, unless you can demonstrate that you will do something beyond the maintenance - that is what can be registered. So, Kyoto intends to get additional projects going, and the money for that is actually quite good - prices of 14 euros per ton of carbon dioxide not emitted have been seen. The price depends on who takes the risk for non-delivery if a project doesn?t deliver the emission reductions on time. The highly energy-intensive South African mining industry is heavily dependent on coal. There should be many opportunities for the pyrometallurgical industry to come up with creative ideas to reduce emissions, and there is now an additional financial incentive to encourage this. Companies can actually get money in return for doing these types of projects, in addition to the contribution to corporate social responsibility.

Jan 1, 2006

By SAIMM SAIMM

To the members of The South African Institute of Mining and Metallurgy We have audited the annual financial statements of The South African Institute of Mining and Metallurgy set out on pages 575 to 586 for the year ended 30 June 2005. These financial statements are the responsibility of the Institute?s Council. Our responsibility is to express an opinion on these financial statements based on our audit. Scope We have conducted our audit in accordance with International Standards on Auditing. Those standards require that we plan and perform the audit to obtain reasonable assurance about whether the financial statements are free of material misstatement. An audit includes: examining, on a test basis, evidence supporting the amounts and disclosures in the financial statements, assessing the accounting principles used and significant estimates made by management, and evaluating the overall financial statement presentation. We believe that our audit provides a reasonable basis for our opinion. Audit opinion In our opinion, the financial statements fairly present, in all material respects, the financial position of the Institute at 30 June 2005 and the results of its operations and cash flows for the year then ended in accordance with South African Statements of Generally Accepted Accounting Practice.

Jan 1, 2005

By G. van den Berg

Anglo Platinum has initiated a project to investigate hydraulic hoisting technology for platinum mines. Hydraulic hoisting systems are categorized as pumping, fluidizing feeder or displacement feeder systems. Positive displacement piston diaphragm pumps are most suitable for pumping systems but they require that the ore be milled prior to pumping. Fluidizing feeders are unsuitable for platinum ore applications due to the large percentage of fines in the ore. The three-chamber pipe displacement feeder has been most widely used for hydraulic hoisting and is the most suitable technology for platinum mines. Eight hydraulic hoisting applications covering pumping and displacement feeder technologies are reviewed. The key factors to be assessed in selecting a hydraulic hoisting system for a platinum mine are discussed. A typical platinum mine configuration will comprise an ore preparation plant (to produce -15mm material through screening or crushing), motive water pump on surface, high pressure water pipeline from surface to underground, ore slurrification plant, three-chamber displacement feeder, and a high pressure slurry pipeline to surface. The water and slurry pipelines are housed in a purpose drilled borehole. A hydraulic hoisting system selection chart is presented. A 150mm diameter system with a 25 MPa pressure rating is required to hoist 60 000 t/m from a depth of 1 200 m. Generic capital and operating costs are presented. A major potential advantage of hydraulic hoisting is that the fines, which contain a high percentage of PGMs, are captured underground and transported directly to the concentrator. Hydraulic hoisting technology is ideally suited to providing supplementary hoisting capacity to existing shafts that have reached the skip hoisting capacity limit. The main advantage for new shafts is the potential to significantly simplify underground rock handling and transportation. All hydraulic hoisting system components are proven. The only risk area is the wear rates of the valves and piping. Current estimates are based on gold ore. The business cases investigated show that the financial viability of the installations are not sensitive to wear rate. Hydraulic hoisting is technically feasible and financially attractive for platinum mines. Keywords: hydraulic hoisting, slurry

Jan 1, 2005

By W. V. Mutler

Present day capital costs are analysed for a 2 000 metric tonne per day (MTPD) pyrite burning plant and compared with a sulphur burning plant of the same capacity. Power, raw material and steam production rates are then compared for the two facility types. A sensitivity analysis is conducted to highlight the impact changing sulphur, pyrite, capital and operating costs have on the relative economics of these feeds. Technological improvement effects on capital cost and energy recovery are also discussed. Besides the steam produced from the burning of pure sulphur or pyrite, plants using an advanced form of heat recovery can produce up to 10 bar steam using heat from acid formation. The quantity depends on the overall water balance which is influenced by such factors as gas strength and product acid concentration. Modern materials and innovative process design also improve the ability to increase the high pressure steam from the roaster boiler in pyrite plants above the traditional 1.25 kg of superheated steam per kg of acid produced.

Jan 1, 2009

By M. I. Brittan

Discussion Dr R. E. Robinson (Fellow): The author must be congratulated on a very meticulous and self-contained piece of work. It is indeed a pleasure to read a paper that is so clearly and systematically laid out, and where the conclusions and the testwork conducted have been so clearly described. The paper is complete in itself, which makes it very difficult for someone who is not intimately involved in the whole Torco project to make any comments on its content. However, two points that, strictly speaking, fall outside the scope of the paper are of considerable interest. The first relates to the particle size of the material treated. In the paper, the testwork is confined to one standard particle size (minus 60 plus 100 mesh). The essential feature of the paper is to indicate that the rate-controlling reaction in the whole segregation process is the rate of reaction of the ore particles with the reducing agent and the hydrogen chloride. One wonders, therefore, to what extent this relatively slow rate of reaction is affected by the particle size of the ore itself. One imagines that the reaction must take place by contact of the hydrogen chloride with the surface of the mineral particles, and it is reasonable to suspect that the rate of diffusion of the copper ions to the surface is a relatively slow process and is thus the limiting factor in this particular rate of reaction. It is possible, for example, that the improvement obtained, when the ore is subjected to reducing conditions before the chlorination, is due to a breakdown in the crystal structure of the original particle. This breakdown is brought about by the reduction and by the consequent increase in surface area available for reaction with hydrogen chloride. Can the author indicate whether any work has been done along these lines, and whether it has been established that the reaction depends on the surface area available? The second point relates to the application of this kinetic study to the actual operation of a Torco reactor. It was once planned to feed the sodium chloride, together with the reducing agent, into the top of the segregation chamber. In the paper, the author mentions that it has now been established that the segregation chamber behaves, to all intents and purposes, as a fluidized bed, and that there is, therefore, a rapid evolution of gas in the lower regions of the chamber, which, it is imagined, displaces the gas phase rapidly. Since the reaction between sodium chloride, water vapour, and the aluminium silicates in the ore is extremely rapid, one wonders how much of the hydrogen chloride produced is removed from the reaction zone before it has had time to react with the copper minerals. The extremely low consumption of sodium chloride (which is a vital feature of the Torco process) must depend on an extremely rapid circulation of the hydrogen chloride gas to all the ore particles in the segregation chamber. One wonders, therefore, if a system for the introduction of the sodium chloride into the bottom regions of the chamber might not result in even greater efficiency in the utilization of sodium chloride. ProC D. D. Howat (Fellow): All of us who have been concerned with the study of chemical reactions at high temperatures are keenly interested in kinetics and are well aware that this is not an easy study experimentally. Dr. Brittan is to be congratulated on the development of neat experimental methods and for his full discussion of the results obtained. Although the segregation process for the extraction of copper from oxide and silicate ores has been known for almost fifty years, the fundamental chemical and physical changes involved have been little understood and the fundamental data are very scanty. The work now in progress at A.A.R.L., together with that sponsored by the Anglo American Corporation in other research institutions throughout the world, is bound to produce new fundamental data and a much more complete understanding of this rather fascinating process. It is already apparent that some of the old and well-worn chemical reactions that were postulated to occur, just cannot take place in the way which was formerly accepted. Dr Brittan&apos;s work, carefully conducted and thoroughly analyzed as it has been, still leaves us with one great outstanding problem. The thermodynamic data and the possible reactions set out in Table I (page 281) of his paper leave us asking, in complete despair, how can copper be converted into a volatile chloride in the presence of HCl, CO and carbon at temperatures about 800°C? The thermodynamics all combine to show that copper should be reduced to metal as the first step in the process. This brings us right up against the second problem. If copper were reduced to the metal how would HCI convert it to the volatile chloride? On top of these problems is the unknown reason for the very high speed of reaction between CO, HCI and the ground copper ore. Still further into the region of the unknown is the reaction by which gaseous hydrochloric acid is produced in the actual process. Perhaps Dr Brittan is feeling grateful that he doesn&apos;t have to try to explain this reaction-at this stage of the research programme at least. The results very clearly show that both CO and HCl gas are essential for rapid production of the volatile copper chloride. Dr Brittan states that 18 minutes were required to attain 83 per cent extraction with HCl gas alone and this was reduced to 4 minutes when CO was

Jan 2, 1970

By M. Tang, X. Tong

"SynopsisSlime coating is one of the most common ways for serpentines to contaminate metallic mineral concentrates during traditional flotation of coarse sulphide particles. This could pose quite a complicated and challenging problem in the case of some types of low-grade and finely disseminated Cu-Ni ores bearing high serpentine contents. This is the case for the copper and nickel sulphides from the Yunnan Mine, China. Previous batch flotation tests of this ore resulted in satisfactory recoveries of 86.92% Cu, 54.92% Ni, and 74.73% Pt+Pd, and concentrate grades of 4.02% Cu, 3.24% Ni, and 76.61 g/t Pt+Pd. However, the MgO content in these concentrates was more than 19%. In the current study, microflotation tests and settling rate tests were introduced to investigate the effects of a combination of strong collectors (a 2:1 weight ratio of butyl xanthate and butyl ammonium dithophosphate) on entrainment of serpentines in metallic mineral concentrate, as well as visual observations of the concentrates in suspension using still photography. All test results indicated the presence of serpentines entrapped in the hydrophobic flocs that resulted from these collectors, even with the use of effective gangue depressants. These strong collectors are used to flocculate the ultra-fine sulphides by forming loose and ‘fluffy’ hydrophobic flocs. However, these hydrophobic flocs may also be able to load or entrap some serpentine slimes into the concentrate, and this entrained serpentine could be harder to remove by using depressants or intensified conditioning than serpentine slime coating on the particle surfaces.IntroductionSerpentine minerals, which have the generalized composition (Mg,Fe)3Si2O5(OH)4, can be easy to crush and grind due to their convoluted and bent layered structures. These serpentines can hinder the enrichment of some metallic minerals and dilute their concentrates by entrainment during flotation. The resulting high pulp viscosities, slime coatings, and high content of dissolved ions adversely affect the flotation recoveries of copper and nickel sulphides, and high levels of MgO entrained in the concentrates lead to heavy penalties from smelters. Since the quality of nickel flotation concentrate depends heavily on its MgO content (generally less than 6–7% for No.1 or 2 grade nickel concentrate according to the standard requirement from National Nonferrous Industry), reducing the entrainment of serpentine during flotation is becoming more and more pressing. Previous research has focused on the collectors and depressants that are commonly used in copper-nickel flotation, as well as their interaction mechanisms."

Jan 1, 2015

By K. -I. Higuchi, M. Mizutani, K. Haga, A. Shibayama, A. Batnasan

In recent years, the quality of iron ores has decreased due to the increase in the content of impurities, namely quartz, aluminum, phosphorus, and sulfur. It indicates that the removal of impurities from iron ore is an extremely challenging task. In particular, the high amount of phosphorus in iron ore could be a problem for energy consumption and CO2 emission in the steelmaking process. To reduce the impact of phosphorus, mainly two approaches: the utilisation of low-phosphorus iron ore, or dephosphorisation of high-phosphorus iron ore using mineral processing technology, have been suggested. In this paper, the removal of phosphorus from a low-grade high-phosphorus iron ore containing 0.1 mass% P and 57.5% Fe by alkaline roasting and wet grinding is discussed. The iron ore consists mainly of hematite, goethite, and quartz. Effects of sodium hydroxide concentration used as an alkaline source, roasting temperature, grinding time, and particle size on the dephosphorisation of highphosphorus iron ore were investigated. The dephosphorisation efficiency, iron loss, and mineralogical components of low-grade high-phosphorus iron ore under various experimental conditions were characterised by using XRD, SEM-EDS, and particle size distribution analysis. Results showed that the phosphorus content was lowered from 0.1 mass% to 0.034 mass%, while the iron grade reached 59.8 mass% from 57.5 mass% by alkaline roasting with 0.05 w/w NaOH at 300 °C for 2 h, followed by wet grinding in an attritor with water for 30 seconds at room temperature. Under the experimental conditions, the dephosphorisation rate was about 54.7%, and the solution obtained by wet grinding contained 22.7 mg/L of phosphorus and 21.5 mg/L of iron. The recovery rate of iron by this process reached 99.9%. Keywords: Iron ore, phosphorus, dephosphoriszation, alkaline roasting, wet grinding

Jan 1, 2020

By S. Kahraman

P-wave velocity measurements were carried out on 22 large limestone blocks in a stone processing plant. In the laboratory, P-wave velocity of core samples was measured to obtain intact rock velocity. The square root of the ratio of the velocity of the large stone block to the velocity of the core specimen was called the velocity ratio index (VRI) and a quality classification for large building stones was suggested. In addition, the applicability of the established classification to stone cutting with gang saws was checked. It was concluded that the quality classification and estimation of slab production efficiency of the building stones can easily be made by ultrasonic measurements. However, the study was carried out only in limestone. Further study must be conducted to check the validity of the obtained results for the other rock types.

Jan 1, 2007

By A. Butler

This paper explains the politics of mine nationalization in contemporary South Africa. Nationalization and privatization typically occur in the oil, mineral resources, and state utilities sectors, and they tend to follow one another in a long-term cyclical pattern. The paper explains why the current international environment has encouraged demands for nationalization and other forms of ?resource nationalism? in the South African minerals sector. It goes on to describe how the resurgence of resource nationalism has been influenced by internal factional politics in the African National Congress. It concludes that nationalization is unlikely to be the outcome of current policy deliberations within the ANC.

Jan 1, 2013