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By Shobhana Dey

A low volatile medium coking coal from Jharia coal field, India was used for this investigation. The proximate analysis of the sample shows that it contains about 25% ash, 21.4% volatile matter and 52.4% fixed carbon. The sizewise ash analysis of -0.5 mm coal indicates that -0.5+0.15 mm fraction contains lower ash (21.9%) than -0.15 mm fraction (29.7%). The flotation characteristics of the -0.5 mm coal were determined by release analysis. The study reveals that recovery of combustible at 15% ash is about 27% and 43% at 17% ash level. The low recovery is due to the presence of high ash in finer fraction. In order to improve the recovery, the -0.5 mm feed was classified into -0.5+0.15 mm and -0.15 mm fractions, and subjected to flotation separately. The recoverable combustible obtained from the release analysis of -0.5+0.15 mm at 15% ash level is 46%. The effect of collector dosage, frother dosage and aeration rate on flotation was studied. The responses of these variables in collectorless flotation of the -0.5+0.15 mm fraction containing low ash were also studied. A lower aeration and frother dosage favored the generation of clean coal with low ash. The ash content in the tailing stream from -0.5+0.15 mm flotation circuit is close to that of the original -0.15 mm fraction. The former is ground to -0.15 mm and the two fractions are processed together. As -0.15 mm fraction contains relatively higher ash, collector aided flotation using sodium silicate was performed to recover the clean coal with 17% ash. The products of the overall flotation circuit having different ash levels were recommended for use in different applications. The reject from -0.15 mm flotation circuit contained 56.4% ash which can be used for fluidized bed combustion (FBC). This eventually leads to complete utilization of coal.

Sep 1, 2012

By R. Tyreman, G. Ritson

"The Red Lake Mine is situated in Balmer Township, Red Lake area in northwestern Ontario. It is an underground mine that has been in production since 1948. The property was originally known as the Dickensen Mine and was shut down in 1996.A feasibility study was carried out in 1998 and it was determined that a new mill should be constructed with a throughput rate of 600 stpd. The life-of-mine average gold grade was projected to be 1.35 oz/t with gold predominantly associated with arsenopyrite and pyrite. Metallurgical testwork indicated that the optimal flowsheet for gold recovery would include an autoclave for oxidation with a carbon circuit for adsorption. To limit capital cost and establish cashflow, the mill was constructed without an autoclave. The recovery by gravity and cyanidation was projected to be 83%. A flotation plant was constructed at the tail end of the circuit to recover the additional gold associated with sulphides from the leach circuit tailings. This material is being stockpiled onsite for future oxidation treatment.Detailed engineering was started in early 1999 and the first gold was poured in August 2000. One feature of the project was the use of Contact Lake equipment from Cameco’s Contact Lake project. The entire mill was purchased and equipment re-used where possible. The capital cost for the process facilities was C$3 1.5 million."

Jan 1, 2001

By J. W. Lyman

The U.S. Bureau of Mines investigated methods of separating valuable rare-earth materials from Fe in neodymium iron boron (NdFeB) magnet scrap. A selective oxidation treatment of the scrap oxidized the rare-earth portion while leaving elemental Fe. Magnetic and leaching procedures were tried for separating the metallic Fe and rare-earth oxides, but the extremely fine grain size of the oxidized scrap prevented recovery by either technique. The best separation of rare earths from bulk NdFeB magnet scrap was obtained by dissolution with H2S04 followed by precipitation of recyclable rare-earth salts. By precipitating neodymium-alkali sulfate double salts as an intermediate that can be converted to a variety of useful products, many materials-handling and economic disadvantages found with direct precipitation with fluoride or oxalate were avoided. Iron was removed from magnet leach solutions by precipitation as a jarosite, eliminating a major disposal problem. Research was also conducted with contaminated mixed SmCo5 and NdFeB swarf. Using a flotation-leaching technique allows the SmCo5 to concentrate in the froth, while the grinding-medium contaminant sinks and is removed as tailing. The NdFeB is dissolved by H2S04 during the process.

Jan 1, 2010

By A Bird

Gravity gold circuits have an important part in reducing operating costs and maximising recovery in modern gold plants where coarse gold is present in the ore body. The design and selection of equipment has a vital part in the overall effectiveness of the gravity gold circuit. There are a number of associated factors that need to be taken into consideration when selecting the most suitable equipment for gravity gold circuits including occupational health and safety, security and labour requirements. This paper is a case study that outlines the validation behind the design, selection of equipment and subsequent benefits of the recent upgrade of the gravity circuit at the Marvel Loch gold mine with the main component to the upgrade being the installation of a Consep Acacia Reactor. Since the upgrade there has been a significant increase in gravity recovery, which has reduced the amount of gold reporting to the leaching circuit therefore reducing reagent consumption and overall operating costs. There has been an increase in overall gold recovery and therefore increased gold production. The benefits to occupational health and safety include the removal of toxic gases being released in the gold room during the previous practice of gold concentrate calcination and the removal of the arsenic and nickel from the gold doré, which poses a risk during refining. Manual handling and security has improved as there is no physical contact with gold concentrates during the automated process using the Acacia reactor. The total time for gold room personnel working in the gold room has also reduced from around six hours per day to less than one hour per day.With the continuing development of low grade gold deposits, there will be an increase in demand for low capital, high recovery gravity gold circuits that can significantly increase overall production while reducing operating costs.

Aug 8, 2011

By J Robinson, L D’Uva

"The Gold Fields group and Gekko Systems have successfully worked together on multiple projects, including at the St Ives, Agnew and Granny Smith operations in Western Australia, and the Damang operations in Ghana, with gravity recovery design, technology installation and optimisation the key driver on each occasion.In 2014, Gold Fields Australia identified an opportunity to challenge the historical thinking at Granny Smith and reassessed the benefit of upfront gravity gold recovery. The historical approach had been to employ gravity on the tailings stream of the carbon-in-pulp (CIP) circuit, recovering sulfide minerals into a concentrate from spiral concentrators for further liberation of gold via a regrind circuit.Gekko was engaged to assess the implementation of a gravity circuit within the grinding area, and to perform a comprehensive gravity recoverable gold test program. Test data was fed into the AMIRA P420 BCC Gravity Model to confirm the gravity recovery potential of the ore, and for process design basis generation. Height, space and capital limitations proved to be a design challenge. However, collaborative learnings throughout the project ensured the design and operation of a reliable and economic gravity gold circuit. Ultimately, the Granny Smith gravity circuit was a success, satisfying all the project objectives and yielding consistent, measurable benefits. On average, the gravity circuit recovers 33 per cent of total gold, resulting in an increase to overall gold recovery by 0.7 per cent.This paper highlights the key project drivers for Gold Fields, and the development of the Granny Smith free gold gravity circuit. It details application of test work and modelling, and the challenges and learnings encountered during the design, installation and commissioning stages. This paper seeks to demonstrate that continuous improvement in a long standing operation can be beneficial to overall circuit flexibility and increased ounce production.Citation:D'Uva, L, Robinson, J and Bell, T, 2016. Defying Gravity - Brownfield Continuous Improvement, in Proceedings 13th AusIMM Mill Operators’ Conference 2016, pp 135–144 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 10, 2016

By J. B. Cabezas

The metallurgical processes that are frequently applied today use strong solutions and are of alkaline, acidic, or organic nature. These play a very important part in recovery of values from many ores. These solutions are of such a chemical nature that they are generally corrosive, flammable, and contaminating. In most cases, they are highly dangerous to the environment. Within the extractive metallurgical industry, it is also very common to have to deal with the handling of contaminating solutions with a high content of cyanide, arsenic, mercury, lead, copper kerosene, acid and many other components whose temporary or permanent confinement within the process must be adequate and absolute.

Jan 1, 1986

By W. Krieglstein

Due to the natural variability of ores, it is necessary to carefully test flotation behavior of flotation cells before designing flotation circuits. For performing quick and versatile tests in the laboratory, Siemens developed a laboratory-size Hybrid Flotation Cell, which has been down-scaled from the 300 m³/h production cell to a 0.6 m³/h flotation cell. This small cell is being used as a test bed for technical and chemical concepts, the improvement of decisive components and the characterization of customer ores. In this study the evaluation results for Andean porphyry copper ores are shown. Total recovery rates of the order of 45 to 50 % have routinely been achieved in a single flotation cycle over a wide range of feed grades. In addition, the Hybrid Flotation Cell has proven to be particularly well suited for the flotation of fine particles below 20 µm down to below 1 µm, at high recovery rates and concentrate grades.

Sep 1, 2012

By A. R. Rule

As part of the Federal effort to help maintain an adequate supply of fertilizer minerals, the Bureau of Mines is conducting research to characterize complex, low-grade western phosphate rock and to develop methods for economic recovery of phosphate. The rock varies from soft, friable, altered phosphatic shale to hard, tightly cemented, unaltered phosphate rock. To supplement current beneficiation by sizing, a flotation procedure was devised that included depression of phosphate minerals with H2SiF6 and anionic flotation of carbonate minerals, followed immediately by cationic flotation of silicate minerals. Application of the procedure to several different ores resulted in phosphate concentrates that ranged from 28 to 31 pct P20S in grade and from 60 to 70 pct in phosphate recovery.

Jan 1, 1977

By Vitaly Kretov, Valery Golikov, Vladimir Ryaboy, Elena Shepeta

"The aim of the study was to develop and implement more efficient dialkyldithiophosphate collectors. A hypothesis has been formulated on the action of dialkyldithiophosphates in the flotation of sulphide ores based on the assessment of their complexing properties and surface activity at the liquid-gas border. Dialkyldithiophosphates form weaker complexes with cations of heavy metals compared to xanthates, and are characterized by a large difference in the strength of their complexes with heavy metal cations. It provides higher selectivity of dialkyldithiophosphates in the separation of non-ferrous metal sulphides and iron sulphides from them. To make a quality assessment of the impact of surface activity of the synthesized collectors on their selectivity of action and collecting power, an assessment of the surface tension of their aqueous solutions has been performed using the stalagmometer method. It is well known that a stronger collector is characterized by a greater surface activity, in contrast to a more selective collector which is characterized by a lower surface activity. Based on studies, new dialkyldithiophosphate collectors BTF 163 and BTF 1552 are suggested for use in the flotation of sulphides to increase the recovery of copper, gold and silver. The collector BTF 163 was introduced at a gold processing plant in Kazakhstan and has been undergoing industrial testing in the sulphide cycle of scheelite-skarn-sulphide ores in a factory in the far east of Russia. The collector BTF 1552, which increases the recovery of silver and gold by 3%, was approved for industrial testing in another far-eastern factory. A modification of the reagent BTF 1552, the new reagent BTF 1522, was introduced in the flotation of copper-zinc ore at a factory in Kazakhstan."

Jan 1, 2014

By A. M. McKim, J. S. Ambler

"The iron ore of the Scully Mine is upgraded from 36 per cent Fe to 66 per cent Fe through single-stage crushing, wet autogenous milling, sizing by vibrating screens and two-stage spiral separation, followed by a process in which the concentrate is dried and then further improved by high-tension separators. The capacity of the concentrator is 17,000 long tons of concentrate per day. Discussion of the operation will recognize the special problems encountered in handling preponderantly fine ore, relate issues of interest regarding wear and review mechanical innovations found to be advantageous at Wabush. The modern trend toward big equipment, thoroughly instrumented, has been the theme of design. Effects on availability must be a consideration in this choice, and sectionalizing of circuits to minimize these consequences is as important as streamlined maintenance."

Jan 1, 1969

By D G. Rantucci

Organic carbon is a major ore contaminant at MMG’s Century mine. This carbonaceous matter is associated with shale hosted sphalerite mineralisation. It occurs as elongated lamellae and as inclusions in porous sphalerite aggregates which are fine grained and difficult to liberate from sphalerite. The carbon prefloat circuit was designed to remove the liberated organic carbon in order to mitigate excessive reagent consumption, contamination of the lead and zinc concentrates and poor flotation performance in the primary zinc circuit. The consequence of organic carbon removal is that sphalerite and galena minerals are lost to the prefloat concentrate. A Jameson Cell was installed in late 2005 as a prefloat cleaner cell. Operational experience showed that the cell became capacity limited during periods of high amounts of organic carbon in the feed and was subsequently bypassed.Reactive style operating strategies were implemented to manage the effects of insufficient organic carbon removal by the prefloat circuit but due to the relatively large circuit residence time, upstream changes meant hours before the downstream circuits were stabilised following a disturbance from excess carbon.This paper describes the modifications and testwork undertaken to enable the Jameson Cell to be operated during periods of high organic carbon in the float feed and the implementation of process control initiatives to proactively remove the required prefloat mass from the float feed to eliminate the reactive style operating strategy.The process control modifications centred on dynamically targeting a prefloat mass recovery determined by an equation relating the carbon to zinc ratio to the carbon prefloat mass as a percentage of the float feed. The equation was developed by analysing historical data. These improvements served to optimise the prefloat circuit and consequently stabilise the downstream lead and zinc circuits.

Aug 8, 2011

By J Seppelt

The Kanmantoo Copper Project is owned and operated by Hillgrove Resources Limited and is situated in the Adelaide Hills region in South Australia. Mining in the area has taken place off and on over the last 140 years to recover metal from both copper and lead-zinc ores. This paper focuses on the commissioning, ramp up and future outlook for the copper concentrator at the Kanmantoo Copper Project with major focus on issues encountered and operating performance achieved to date.CITATION:Seppelt, J, 2012. Commissioning of the Kanmantoo Copper Concentrator, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 127-136 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

By Robert A. Jankowski

Through laboratory and underground studies, the Bureau of Mines has evaluated the use of passive barriers (dust shields) to help confine dust generated by longwall shearers to the face area. Laboratory studies focused first on selection of the optimum barrier design, which proved to be a simple, gobside barrier (made of conveyor belting), coupled with a headgate splitter arm, which was found necessary to prevent the dust cloud from entering the walkway at the headgate end of the shearer when cutting tail to head. Further laboratory testing was done to determine whether using passive barriers in combination with different water-spray dust-reduction methods, including the shearer-clearer system developed by the Bureau, would improve dust reduction over that from using the water sprays alone. It was found that the combination of gobs ide passive barriers plus the full shearer-clearer system was most effective, and surmised that this would be especially useful in thicker seams. However, underground testing revealed that the shearer-clearer system alone was 35 pct more effective in reducing dust contamination than when used with the gobside passive barriers, and far more effective than using passive barriers with a conventional water spray system. For such an effective system as the shearer-clearer; the reduction in avail-able space over the top of the machine actually impeded system performance; this was shom1 in both eastern and western mines. But when used with comparatively inefficient water spray systems, the passive barriers, used in combination with a headgate splitter arm, provide consider-able help in reducing dust levels in the walkway.

Jan 1, 1986

By ERNEST W. ELLIS

MORE or less contradictory findings as to the most satisfactory feed for concentration tables of the Wilfley type is shown by the diversity of opinion among experimenters. Prof. R. H. Richards,l as a result of his work, favors classification. On the other hand the work of Messrs. Cox, Gibbins, and Porter,' of McGill University, points in general to the desirability of screen-sized feed The ore used was obtained from the West mill of the Bunker Hill and Sullivan Mining and Concentrating Co., Kellogg, Idaho, and consisted of argentiferous galena in a gangue of quartzite and siderite. The natural feed was taken from the launder supplying the Fahrenwald classifier. Classified feed was collected from each of the ten spigots of this same classifier. From the natural feed the screen-sized feed was pre- pared in the laboratory. All three types of table feed were therefore derived from the actual mill circuit of a large plant.

Jan 1, 1926

By R. L. Keiski, H. Miettunen

"The serpentinite massifs of Hitura are located in the municipality of Nivala in Central Finland. The most important sulfide minerals in the ore are pentlandite, pyrrhotite and chalcopyrite. The ore body consists of amphibolite rocks in contact with serpentinite against mica gneiss. The mean Ni grade of the ore is about 0.7%, and the highest Ni grades have been encountered in close contact. In general, the inner parts of the ore are strongly serpentinized ultramafite and some sulfides have been converted to mackinawite and valleriite. The degree of alteration varies considerably within the ore body and, correspondingly, the grain size of the sulfides varies as well. In this study, flotation tests were done with samples from different zones of the ore body. The flotation test results on the bench scale and plant scale are discussed in terms of the mineralogical features of the ore. Ores in the North Hitura zone differ from those in the South Hitura zone.IntroductionThe Hitura nickel mine, like the entire Nivala area, is part of the Bothnian Svecokarelian schist zone, where numerous Ni-Cu boulders have been discovered since the early 1930s. The Nivala area has therefore been the subject of exploration activities by the Geological Survey of Finland (GSF) and mining companies. The first period of exploration started in 1936, and the Makola nickel deposit, located 3 km south of the Hitura mine, was discovered by GSF in 1937. About 415,000 tonnes of ore were mined at Makola between 1941 and 1954. In 1937, during the exploration and development stage of the Makola mine, boulder tracing and a geophysical survey revealed a significant ultramafic body with some lowgrade Ni-Cu occurrences in the South Hitura zone.In 1961, GSF continued the exploration program in the area and discovered the North Hitura ultramafic complex and its allied Ni-Cu occurrences. In 1964, Outokumpu Oy bought the property and, after underground exploration, comprehensive mineralogical studies, process research and a feasibility study, production started in 1970 (Isohanni et al., 1985)."

Jan 1, 2015

By Xiaoping Xu

An albitization granite type tantalum-niobium ore from Jiangxi Province, China was taken as the research object in this paper. The main valuable minerals in this ore are tantalite-columbite and uranmicrolite, and the minerals that can be comprehensively recovered include cassiterite, sulfide minerals, topaz, zinnwaldite, and quartz-feldspar. Based on the research results of process mineralogy, the roughing process of gravity separation for coarse particles combined with flotation for fine particles is proposed. The combined cleaning process of flotation-gravity-magnetic-electric-hydrometallurgical separation is adopted for the rough concentrate of gravity separation while the combined cleaning process of gravity-flotation- hydrometallurgical separation is adopted for the rough concentrate of flotation. High intensity magnetic separation is used to separate and recover zinnwaldite and quartz-feldspar from gravity tailings. By mineral processing for the crude ore containing (TaNb)cassiterite concentrate is 0.0003% with a grade of Sn 63.22% and a recovery of Sn 47.42% while the yield of sulfide concentrate is 0.0783% with a grade of Pb 33.15% and Zn 27.09% which can be separated to get lead and zinc concentrate respectively. The yield of topaz concentrate is 0.1545% and the topaz mainly distributed in the size range of -0.4+0.074mm, which can be made into abrasive products of different particle sizes after classification. The yield of zinnwaldite concentrate is 16.91% with a grade of Liwhiteness of 75%, which is a glass-ceramic material of good quality. 2O5 0.0346%, Sn 0.04%, Pb 0.10% and Zn 0.33%, the tantalum-niobium concentrate with a yield of 0.00278%,a grade of (TaNb)2O5 53.61% and a recovery of (TaNb)2O5 44.49% is obtained. And the yield of 2O 1.34% and Rb2O 0.40%. The yield of quartz-feldspar concentrate is 30.42% with a grade of K2O+Na2O 8.98% and Fe2O3 0.11% and a natural All of tantalum, niobium, cassiterite, topaz, zinnwaldite, quartz, and feldspar are recovered effectively. Keywords: tantalum-niobium ore, gravity separation for coarse particles, flotation for fine particles, comprehensive recovery

Sep 1, 2012

Research 83 highlights significant studies conducted during the past fiscal year at Bureau facilities and offices. Readers who want more information about the projects and studies summarized here can contact the researchers whose names and locations appear with each summary, General information on Bureau of Mines programs is available from the Office of Technical Information, 2401 E Street, NW., Washington, DC 20241. A listing of principal officials, a location map of field facilities, and a selected bibliography of publications and patents issued on Bureau developments during the past 3 years are provided.

Jan 1, 1983

By Ralph H. Light

The recent surge in demand for uranium in the United States, accompanied by a significant increase in price, has resulted in exploration, mining, and milling of lower grade ores than have been exploited in the past. To sustain production levels of uranium with lower grades of feed requires larger tonnage mill operations. Also, because of recent environmental regulations, mill operations must be prepared to properly contain large tonnages of tailings, and recycle, treat and discharge, or evaporate mill effluents. Thus, the advent of lower grade ores, larger tonnage mills, and environmental considerations has made the cost of milling a critical factor in today's uranium production economics in the United States. Three basic milling methods are employed for extracting uranium from ores: 1. Acid leach, countercurrent decantation, and solvent extraction. 2. Acid leach, resin-in-pulp, and Eluex. 3. Alkaline leach, solid-liquor separation, and precipitation. Vanadium often occurs as a valuable by-product in ore deposits located in the Uravan mineral belt in Western Colorado. Also, lesser amounts of vanadium may occur in ores from South Dakota, New Mexico, and Utah. Vanadium is usually extracted in the acid leach solution with uranium and recovered by solvent extraction after recovery of the uranium. By-product recovery of uranium from phosphoric acid and porphyry copper dump leach solutions is now receiving considerable attention. Pilot plant studies are under way and construction of commercial plants is in progress.

Jan 1, 1977

By V A. Chanturia, T N. Matveeva, V V. Getman, A Yu Karkeshkina

The development and application of new effective environmentally friendly reagents for the flotation of thin and ultrathin mineral components is an urgent task. The analysis of the nomenclature of chemicals, including flotation reagents and flocculants, showed that water-soluble polymers (WSP) might be used in flotation separation of fine grinding products containing non-ferrous and precious metals. Water-soluble polymers that are responsive to temperature changes are the most attractive as reagents during flotation. These are the thermo-responsive polymers or thermomorphic polymers. The presence of nitrogen and oxygen atoms in the polymer molecule contributes to the formation of a coordination bond with non-ferrous and precious metals. A research conducted at the ICEMR RAS within the framework of the scientific school of Academician of the Russian Academy of Sciences Chanturia helped to develop a method of flotation of finely ground mineral particles containing valuable components using modified thermomorphic polymers. The paper focuses on the experimental study of poly (vinyl caprolactam) (PVC) as a flotation reagent for gold recovery from finely ground mineral products. PVC is a ready-made water-soluble thermosensitive polymer that does not require complex synthesis. PVC has a high ability to form complexes in an aqueous environment when heated. According to experimental data, the polymer interacts with gold nanoparticles, contributing to the extraction of gold from fine mineral products and the tailings that makes it perspective as an additional flotation agent for the processing of gold-containing ores.

Aug 24, 2022

By S. F. Diehl

Suites of monomineralic sulfide and carbonate minerals, as well as mine-waste composite samples, were analyzed using chemical and microanalytical techniques to understand behavior and bioavailability of mine waste elements such as Cu, Fe, Pb, or Zn. Petrographic, scanning electron microscopy (SEM), and laser-ablation inductively-coupled mass spectrometry (LA-ICP-MS) studies of mineralogical residence sites of trace elements are important because trace elements affect the chemical and physical behavior of minerals. Leachate solutions show that chalcopyrite, bornite, azurite, malachite, and galena are host to the trace elements Ag, Al, As, Cd, Co, Cr, Mn, Ni, Pb, and Zn. SEM studies reveal whether these trace elements occur within mineral inclusions or are substituted within lattice structures. SEM analysis reveals mineral phases not identified by X-ray diffraction (XRD), such as inclusions of sphalerite and cobaltite(?) within chalcopyrite.

Jan 1, 2010