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By Peter Waggitt

Northern Australia has been home to uranium mining for over 70 years, more or less continuously. In the early days, post World War 2, there was little effort made to remediate workings at the end of operations. The three “campaigns” of uranium mining (1951-55; 1960-70; and 1979-present) are all now in various stages of remediation. This includes not only the large legacy site at Rum Jungle but also smaller fields like the South Alligator Valley, as well as contemporary works at Nabarlek and the ongoing closure at Ranger– Australia’s longest operating uranium mine. This paper briefly describes some of the past remediation activities before providing an update on the Ranger closure programme and future timetable.

Jan 1, 2020

By G. Deschênes

Located in North West Tanzania, Buzwagi is an underground narrow-vein gold mine owned by Barrick Gold. The plant was commissioned in 2000. The reserves are estimated to be more than 25 years. The ore, containing gold, silver and copper mineralization in sulphides, is processed at 3,300 tpd. This paper presents the results of an investigation to determine the performance of an ammonia-cyanide leaching approach to recover gold from two samples. Coarse gold was recovered by gravity prior to leaching. The gravity tailings had the following composition; 1.97 g/t Au and 0.31% copper for the medium cyanide soluble copper (MCSC) sample and 2.80 g/t Au and 0.53% Cu for the high cyanide soluble copper (HCSC) sample. Experiments were conducted for 72 hours in stirred tank reactors with monitoring and control of parameters. The straight cyanidation produced gold extractions of 76.1% for the MCSC sample and 84.6% for the HCSC sample. Associated cyanide additions were 6.1 kg/t and 11.5 kg/t respectively. The gold extraction of the MCSC sample was improved to 88.8% using 350 ppm NaCN and 700 ppm ammonia. The associated cyanide and ammonia additions were 2.5 kg/t and 1.7 kg/t respectively. For the HCSC sample, 82.5% gold extraction was obtained using 500 ppm NaCN and 600 ppm ammonia. The cyanide addition and ammonia additions were reduced to 3.8 kg/t and 1.0 kg/t respectively. No significant decrease in gold extraction was noticed when the cyanide and ammonia were reduced to 150 ppm and 500 ppm respectively. In the presence of ammonia, a significant amount of cyanide was oxidized to cyanate by the cupric ion. In the absence of ammonia, the formation of copper cyanide was the predominant reaction responsible for consumption of cyanide.

Jan 1, 2012

By Yakup Cebeci, Leyla Kalender, Ceylan Yarluğkal Altınışık, Hikmet Sis

The abstract highlights the challenges in concentrating complex and non-sulfide lead-zinc ores and emphasizes the importance of froth flotation for their beneficiation. The study focuses on using process mineralogy to understand the flotation behavior of these ores from the Görgü region. The ore samples were characterized using various techniques, revealing oxidation and carbonation of primary minerals. The flotation results were interpreted in light of the mineralogical data, explaining the poor outcomes due to the complex nature of the ore, poor liberation, and lack of hydrophobicity.

Mar 22, 2022

During operations, mining and metallurgical processes may generate effluents such as tailings water, acid mine drainage (AMD) and seepage and process acid streams. Depending on the type of ore and the metallurgical process, these effluents may contain one or more toxic compounds (eg acidity or alkalinity, cyanide, ammonia and/or nitrate, heavy metals, total suspended solids (TSS), sulfate) in elevated concentrations requiring treatment before their discharge to the environment or recycle/reuse in the process. Natural oxidation of sulfide minerals present in mining wastes (eg tailings and waste rock) at mining sites may generate AMD, which is characterised as a low pH, high acidity effluent containing heavy metals and sulfate. If generation of AMD cannot be controlled and/or prevented, AMD needs to be collected and treated for neutralisation of acidity and reduction of metals and TSS to meet regulated water quality standards. It is important to note that some of these compounds may persist in run-off from mine sites, even after the mine and processing facilities are decommissioned. Treatment of mining and metallurgical process effluents can be accomplished by means of various physical, chemical and/or biological methods. The mode of process application may vary from the use of either specifically designed, controlled and automated facilities or passive systems. Lime neutralisation and precipitation is the most common method used in the mining industry to treat AMD. To reduce the problems associated with disposal and long-term storage of the resulting sludge, the use of a high-density sludge process (HDS) has become a preferred option. The use of other chemical reagents, waste or by-products from other industries, and biological sulfate reduction methods can also be considered as viable options for site-specific situations. Recently, several passive treatment systems have been designed and successfully operated, even in the cold North American climate. This paper will discuss available options and provide insights for selecting a suitable method for a given situation using case studies from projects conducted by Golder Associates around the world over the last decade.

Jan 1, 2006

By Chas. F. Jackson

This circular is an abstract of a manuscript-completed in August 1932, which cannot be published in its entirety due to lack of printing funds. It was originally planned to publish the manuscript as one of a series of bulletins dealing with the mining and milling of ores of the principal metals. Because of the common association of lead and zinc minerals in many ore deposits, and because a considerable proportion of the zinc produced in this country is obtained as a by-product of lead mining or vice versa, it seemed necessary to treat the mining and milling of lead and zinc ores in the same volume.

Jan 1, 1934

By Shounak Banerjee, Suresh Babu Dasari, Dharmendr Kumar, Shashank Pardhikar, Beena Rai, Venugopal Tammishetti

Recent increase in steel demand has led to the increased usage of high grade iron ores. During the mining, crushing and washing of ore, huge quantities of fines and slimes (<150 µm) are generated, majority of which is dumped into tailing ponds owing to their high alumina contents (~10-15%). Due to rapid depletion of high grade iron ore, it is the need of the hour to utilize low grade iron ore fines and slimes. These fines and slimes contain high iron values assaying 45-60% Fe. Dumping these valuable resources in ponds locks enormous land space and causes environmental hazards. The constituent minerals present in Indian iron ores comprise of hematite, goethite, gibbsite, kaolinite and quartz. Our research is focused on beneficiation of iron ore slimes through selective flocculation. We use density functional theory simulations to screen the reagents and perform experiments with screened reagents to determine the experimental conditions for effective performance of the reagents. Our studies revealed that reagents like starch, guar gum and carboxymethyl cellulose exhibit excellent selectivity in high alkaline pH range. However, the higher cost of alkali makes the process economically less attractive. In the present study, a commercial reagent from NALCO Water (Hi-Tex 83570) has been explored as selective flocculant for beneficiation of iron ore slimes in wide pH range (4-11). However, the reagent was also found to be the most selective in alkaline pH range only. Full factorial design of experiments was done considering three parameters namely, pulp density, flocculant dosage and settling time. Regression models were developed for yield, iron grade, alumina grade and iron recovery in the concentrate, using experimental results. Further parameter optimization was done to achieve high Fe recoveries and low alumina recoveries. The optimal parameters thus arrived at were validated with the experiments. We were able to achieve a concentrate grade of 60.0% Fe and 5.1% Al2O3 with 68.8% Fe recovery from a feed containing of 53.1% Fe and 8.9% Al2O3. The performance of flocculant Hi-Tex 83570 was also compared with reagents starch and guar gum.

Jan 1, 2018

A processing scheme is presented to use one single chemical that has multiple functions to re-duce the use of chemicals and achieve more efficient mineral flotation and solids dewatering. Sti mulisensitive polymers which display hydrophilic/hydrophobic transitions in response to changes in temperature or pH have been found to be useful as such multiple function reagents. They can cause the mineral particles to be either hydrophilic or hydrophobic at one stimuli condition and the opposite at the reverse stimuli condition. Therefore, both particle surface wettability and inter-particle interaction forces are reversibly controllable. When the surface is hydrophilic, particle dispersion is achieved by repulsive inter-particle forces whereas when the surface is hydrophobic, particle aggregation is induced by inter-particle hydrophobic attractive forces. In addition, the hydrophobic surface condition allows for the attachment of particles to bubbles. Flotation and solid settling tests have been conducted with a silica suspension treated with the temperature-sensitive polymer poly (N-isopropyl acrylamide) (PNIPAM). The results indicated that PNIPAM can be used to induce a transition between hydrophilic and hydrophobic surface character at temperature lower or higher than its critical solution temperature (CST around 32oC). Thus, both effective silica dispersion or hydrophobic aggregation and flotation without adding any additional collector have been demonstrated. In tailings treatment, rapid silica settling was obtained with hydrophobic aggregation at temperature above the CST and further sediment consolidation (and water release) occurred at temperature below the CST. Contact angle measurements confirmed the hydrophilic/hydrophobic transition induced by PNIPAM.

Jan 1, 2014

By Mingfei He, Xianyang Qiu, Hongling Wang, Yude Gao, Qingbo Meng, Xiaoping Xu, Hao Bu, Guosheng Wang

"By the “HGMS-gravity separation-Flotation” combined technology, a Ta-Nb-Sn concentrates with the grades of Ta2O5 10.00%, Nb2O5 9.13% and Sn 42.62% at the recovery of 44.73% , 9.13% and 14.62% respectively , a Li concentrate with Li2O 3.11% at the recovery of 51.83% and a feldspar product are obtained. All the value elements are recovered. INTRODUCTION Tantalum and niobium widely used in electronics, atomic energy, aerospace and chemical industries with the advantages of high melting point, good ductility, low vapor pressure, strong corrosion resistance and high thermal conductivity are important strategic resources which occupy a special position in rare metals (Crockettand, 1993; GAO and Zou, 2004; LI and HUANG, 2012). China's tantalum and niobium resources are poor, and most of them are low grade, fine grain size, complex composition, and polymetallic associated with refractory mineral resources (GAO and Qiu,2013; Liang and Li, 2016). The Yifeng Lion Ore containing lithium and other rare metals (tantalum, niobium and tin) is an important lithium polymetallic deposit in China Jangxi province. While extensive previous studies have settled on the recovery of lithium minerals by flotation only for the extremely low grade of Ta, Nb and Sn. The regular gravity concentration is used, and the recovery of tantalum niobium is low. This paper sets out to investigate a combined flowsheet of “high gradient magnetic separation (HGMS),-gravity separation-flotation” to effectively comprehensively recover the Li, Ta, Nb, Sn and the feldspar resources in the Yifeng Lion Ore.ORE PROPERTIES Chemical analysis The run of mine comes from Jiangxi Yifeng Lion ore. The main elements of the ore chemical analysis results shows in Table 1. From Table 1, the grades of lithium and rubidium are relatively high, but the grades of tantalum and niobium are lower than the industrial grade; feldspar with high content of potassium and sodium and low content of iron can be used as high-quality ceramic raw materials."

Jan 1, 2018

By F. C. Araújo, T. Tavares, J. Leite

"This work presents the small-scale mining cluster of kaolin in Rio Grande do Norte (RN) and Paraiba (PB), both Brazilian states, including the following information: number of companies, process flowcharts, production numbers, mineralogical (XRD) and chemical (XRF) composition, granulometric distribution (screen and laser granulometer) and image analysis (SEM). Some social indicators in the region of the RN and PB States are presented in association with these data. The number of companies and their flowsheet were obtained through in situ visits and sample collection for technological characterization. The comparative study of the different flowsheets for kaolin production in the world was done using the databases and presenting the state of the art of kaolin processing. The results indicated that the mean feed typically contains 35-39% kaolin with quartz, muscovite and kaolinite as main mineralogical components and SiO2 (65.83%), Al2O3 (21.19%), K2O (3.9%), and Fe2O3 (0,63%) as main chemical constituents. The products of mineral processing are twofold, one between 74 µm and 37 µm, and the other smaller than 37 µm. The wastes of operation had major quantities of kaolin. The average Human Development Index (HDI) of the cluster region is approximately of 0.61, which shows an association between mineral processing plant and social indicators. The kaolin recovery, however, is between 30-55% for the three most common types of flowsheets that are currently used in the small kaolin mining cluster, indicating a need for technological improvement to achieve regional development. These data could be used incrementally to adapt the process plant currently used in the small-scale mining cluster of kaolin."

Jan 1, 2014

By W. S. Hargratt

THE present era of exploitation of the Little Long Lac and Sturgeon River areas commenced in 1925, when Beardmore Gold Mines started development of what is now the property of Northern Empire Mines. A shaft was sunk in 1927-28 and a very limited amount of underground work was carried out, following which the property was shut down. It remained idle until 1932, when Empire Star Mines, a Delaware Company, optioned the property. This Company formed the Northern Empire Mines Company, Limited, to operate the mine, and immediately instigated an aggressive programme of surface and underground development. By the summer of 1933, sufficient ore was indicated to justify the installation of a small mill. The decision was also reached to bring in electric power. A power-line was built from Hydro, the Ontario Power Commission&apos;s development on the Nipigon river, to the mine, a distance of approximately 50 miles. This work was entirely financed by Northern Empire Mines. Mill construction was commenced in October, 1933, and completed by March, 1934, thus making this mine the first producer in this district.

Jan 1, 1940

By R. L. Grayson, S. S. Peng

"Abstract. Sampling of respirable dust and analyses of samples have been partially completed for a longwall panel. Physical and compositional properties which have been characterized include:l. Respirable mass concentration, 2. Mineral composition,3. Particle size distribution by mineral,4. Particle shape distribution by mineral,5. Particle angularity distribution by mineral, and6. Respirable size distribution.Analytical techniques used in the characterization of samples are described, and results are reported with emphasis placed on relationships between the various characteristics and worker locations. Details of existing mining conditions, equipment arrangements, dust control methods and operating system parameters are presented.INTRODUCTIONThrough the work of many researchers worldwide, the health of coal miners in modern societies is now protected from dangerous exposure to respirable coal mine dust. A vast amount of research culminated in the exposure-response model of Jacobsen, et al. (1971) whi3h has been used to establish the 2mg/m exposure limit in the United States. According to the model, exposure to respirable dust for 35 years at this level or less would prevent the occurrence of a clinically significant case of coal workers' pneumoconiosis (CWP). The exposure limit, of course, must be adjusted whenever greater than five percent free silica exists in the respirable dust.The Jacobsen model was built around a mass concentration exposure, which showed the highest correlation with incidence and progression of CWP. As Morgan; et al. (1973) noted, howev.er, the quantity of respirable dust inhaled is not the only significant factor. It is noted that although mass of respirable dust correlates highest with the incidence of CWP, it is not a causal factor. Indeed, the insidious processes causing the disease occur on the cell-particle scale, requiring microscopic investigation."

Mar 1, 1989

By The Staff

"THE property of Red Mountain Mines Limited is located 11/2 miles north and west of the town of Rossland, in southern British Columbia. The mine is situated on the flank of Red mountain, famous in pioneer days as the site of a number of profitable gold-copper producers. Mining and milling commenced in May, 1966, and by the first of July the concentrator was treating close to its average design capacity of 400 tons per day. The following description of the property and its operations is divided into three parts, dealing respectively with geology, mining and milling.GeologyAll the known ore reserves of Red Mountain Mines occur on the Coxey claim, on the west slope of Red mountain. Although there are indications of molybdenite mineralization elsewhere on the Company's property, these have not been developed.The Coxey and adjoining areas are underlain by the Mount Roberts formation of the Rossland volca-nics. This formation consists of an interbedded series of volcanic and sedimentary rocks, mostly brecciated and silicified near the ore zone, and all intruded by igneous rocks which occur in various forms, and vary in composition from basic to acid. The country rock in the vicinity of the mineralization seems to be mostly a volcanic, approximately andesitic in composition, and possibly a surface flow. There is much banding, some of which is undoubtedly due to tuff and agglomerate layers, and some to narrow, siliceous, sedimentary horizons such as the surface exposure of the main ""A"" orebody, much of which was a thin capping of greyish banded rock that looked like quartzite. The drill holes also intersected occasional layers of black, argillaceous quartzite, sometimes altered to a greenish colour. The whole assemblage has been silicified with varied intensity, sometimes to such an extent as to have obscured or obliterated the original characteristics. The mineralized zone has also been strongly brecciated, the fragments varying in size from less than an inch up to large masses, best described as a mega-breccia. Anomalous results obtained in drilling, such as sedimentary sections that do not appear in adjoining holes, may sometimes be due to the presence of these large blocks, no longer in their original position. The whole volcanic unit is thus very complex and confusing. For practical purposes, the ore may be said to occur in a greenstone formation with minor sedimentary interbeds and considerable later igneous material."

Jan 1, 1967

By Walter R. Storms, Allan B. Bowman

This paper , which describes mining methods and practices at the Mineral Hill copper mine , is one of a series being prepared by the Division of Minerals of the Bureau of Mines on mining methods , practices , and costs in various mining districts. of the United States . The Mineral Hill mine is situated on the northeastern piedmont of the Sierrita Mountains at an altitude of 3,650 feet in sec . 35 , T. 16 S. , R. 12 E. , and sec . 2 , T. 17 S. , R. 12 E. , Pima mining district , Pima County , Ariz . ( fig . 1 ) . The claims (figs . 2 and 3) are in a gently rolling area with only scant desert vegetation . The average annual precipitation is about 11 inches . The nearest surface water is the Santa Cruz River , an intermittent stream 8 miles east of the mine . The property is owned by the Banner Mining Co. , a Nevada corporation . This report presents a brief history of the mine , describes the geology of the area and the ore deposits , outlines methods of prospecting and exploration , and gives methods of sampling and the estimation of ore reserves and values . Development and stoping methods are explained , and methods of underground transportation , ventilation , mine drainage , and safety measures are described . The concluding section of the paper is a section on milling . The Mineral Hill mine is an old mine that was reopened in 1951 after having been closed for 30 years .

May 1, 1957

By M. Guzman, M. Romero, S. Bravo-Thais, M. Flores

The phytoremediation method is proposed to recover valuable metals from polluted soils. In this sense, the objective of this research is to determine if it is possible to use native plants in the recovery of precious metals from abandoned mining liabilities. The tailings samples containing Au and Ag were acquired from the northern region of Peru. Two native plant species, Stipa Ichu and Cortaderia Nitida, were used. Results show that Au was recovery around 19% (1.52 mg/Kg) and 25.43% (1.78 mg/Kg) by using Stipa Ichu and Cortadeira Nitida, respectively. Besides, Ag had a better recovery in the Stipa Ichu reaching a value of 6.83% (0.628 mg/Kg); in contrast, Cortadeira Nitida had an efficiency of 3.6% (0.814 mg/Kg). Bioaccumulation, bioconcentration and translocation factors were calculated by determining the concentration of precious metals in the roots and aerial parts of each plant. Results show that Stipa Ichu had a greater capacity of hyperaccumulated Ag compared to Au; while Cortaderia Nitida tended to have a better performance for Au. In conclusion, both plants are recommended for phytoremediation of mining liabilities.

Feb 1, 2023

By Max J. Spendlove

The Federal Bureau of Mines has been a leader in resource recovery research since it was founded in 1910. Until the early 19601s, most of the projects were concerned primarily with the task of extending and improving conventional technology oriented to secondary metals and minerals upgrading and processing for marketable products of highest value. By the mid-1960&apos;s with the passage of the Solid Waste Disposal Act of 1965, all resource recovery research programs were increased in scope and intensity as an effort to solve the growing number of environmental problems and the rapid depletion of domestic energy and mineral resources. Many new materials were investigated including urban refuse, manufacturing, industrial wastes, and stockyard wastes. The objective was to promote maximum recycling as the most practical answer to problems of waste disposal, pollution control, and the short supply of minerals and energy. The new research projects ranged from the conversion of animal manure and other organics to oil, to the recovery of precious and common metals from demolished aircraft and electronic scrap. This profile includes brief descriptions of many selected research projects that typify, the total spectrum of Bureau of Mines research in the important field of resource recovery.

Jan 1, 1977

By Arthur L. Crawford, R. E. Head, Fred E. Thackwell, A. Lee Christensen

"The data in this report are the result of a detailed, microscopic study of composite samples of mill products collected over a 14 day period. The samples were collected and assembled under the direction of W. R. Lindsay, superintendent of the Anyox plant, and are fairly representative of the heads, middlings, tailings, and products peculiar to the scheme of concentration practiced at Anyox in 1930.The work was done in the Cooperative Microscopic Laboratory at Salt Lake City, Utah. This laboratory was equipped by the mine and mill operators of the Salt Lake Valley and functioned under a cooperative agreement between the United States Bureau of Mines the University of Utah. This is the fourth of a series of papers 6/ dealing with a microscopic study of mill products."

Oct 1, 1935

By Kwadwo O. Adutwum, George M. Potter, Todd J. Harvey

Ashanti Goldfields Company (AGC) is the largest gold mining company in Western Africa. Their jewel operation is a highly inte­grated set of mines and plants situated in Obuasi, in the Ashanti region of Ghana. This operation is responsible for mining 5.5 mil­lion tonnes of ore at 4.6 git annually from a series of open pits and underground mines. The sulphide ore is highly refractory with untreated gold recoveries in the range of 30-40%. The milling facilities are quite varied and include an oxide treatment plant (OTP), a heap leach, a sulphide roaster circuit (PTP), a tailings treahnent plant (TIP) and the most recent, a sulphide treatment plant (STP). The sulphide treatment plant, the world's largest stirred tank biooxidation circuit, was commissioned in early 1994. This plant treats approximately 800 tonnes of refractory gold con­centrate per day with overall recoveries of75-85o/o(ll, Biooxidation was selected for its ease of operation and less exotic maintenance requirements as well as its ability to achieve comparable gold recoveries. Unfortunately, despite stirred tank biooxidation being one of the lower cost methods of treating refractory ores, it is still too costly for low grade refractory materials. GeoBiotics, Inc. has developed a proprietary technology that allows the economic treatment of lower grade refractory gold ores through heap biooxidation. AGC has a significant resource of low grade refractory sulphide ore. This ore is too low grade to warrant BIOX™ sulphide treatment and its oxide plant recov­ery is low at 35-45%. GeoBiotics and AGC have completed pre­liminary testing and economic analysis of heap biooxidation for this material. Additionally, the feasibility of producing a gravity concentrate from the tailings of the sulphide treatment plant has been explored. This concentrate would be coated onto the low grade sulphide and biooxidized concurrently. Gold recoveries for the biooxidized low grade sulphide have increased to 75% and the gravity recovery from the STP tailings has achieved 80% uti­lizing a Falcon concentrator. The preliminary economics of the proposed full scale plant are very favourable with a pre-tax IRR of 233% and a payback period of 6 months. This paper details the biooxidation testing, flow sheet devel­opment and preliminary economics of employing heap biooxi­dation for low grade refractory ore treatment at Ashanti Goldfields' Obuasi operation.

Jan 1, 1998

By M. Albertson

At Cobalt, Ontario, Canada, a lake was drained to facilitate mining, by the Mining Corpn. of Canada, during the spring and early summer of 1915. Previous to pumping out the water, great quantities of sands and slimes from concentrating plants had been discharged into the lake and during and after the lake&apos;s drainage, its basin was a receptacle for tailing products. As the writer was at work along the shore line as the lake was being drained, he had a good opportunity to observe the changes that took place as the water was withdrawn. Some adjustments between the incoming sands and the mud in the lake had taken place before pumping was commenced. One of the most interesting results was the appearance of a small dome in a path the writer traversed twice a day for several months; he remembers distinctly the difficulty of crossing this. Cobalt Lake owed its existence to the gouging out of a rock basin by glaciation. The long axis of the basin closely follows the strike of a thrust fault of about 500 ft. (160 m.) vertical displacement. The lake was originally shaped somewhat as shown. The length was about 3000 ft. (914.4 m.), the width at the lower lobe was about 1000 ft., and the width at the narrows probably 400 to 500 ft. The original depth of water varied from 20 to 30 ft. (6.1 to 9.1 m.), near where the island later was formed, to 60 to 70 ft. (23.6 to 27.6 m.) in the widest part of the lower lobe. At the narrows, the depth was 30 to 40 ft. Above the bed rock was sand, with boulders near the bottom, and mud. During the building of the railroad in 1903, considerable filling was done along the right of way. With the commencement of mining operations, about 1905, waste rock and mill tailings were dumped into the lake. One of the mining companies operated a hydraulic giant to remove the glacial debris from several hundred acres of rock surface; much of the sand and clay from this operation was deposited in the lake. About 1,500,000 tons of mill tailings, composed of sands and slimes, were discharged into the lake&apos;s waters previous to 1915. Most of this material

Jan 1, 1921

By H. C. L. Reemeyer

One of the fundamental trade-offs for copper concentrators is between grinding circuit product size and copper recovery. The capital and operating cost savings and throughput increases for a coarser grind should be evaluated against potential recovery reductions. Other benefits of coarser grinds, such as GHG emissions reductions and favourable effects on tailings and water management should also be considered. This paper assesses the formation of laboratory and pilot test programs, design of process flowsheets, equipment selection, and evaluation methods to determine the most appropriate and economical grind size for a given deposit. The paper will include an evaluation of case studies. INTRODUCTION The majority of the world’s mined copper comes from sulphide deposits, which are mostly processed using crushing, grinding, flotation and dewatering flowsheets to produce copper-bearing concentrates for smelting. Ores must be ground to a sufficiently fine size to achieve an acceptable level of liberation to allow both high recovery of copper and valuable by-product minerals while efficiently rejecting gangue to tailings. Therefore, a key design parameter in process plant design is the grind size of the flotation feed. Copper concentrators use a variety of grind size targets that depend on various factors, including ore mineralogy and the grain sizes of valuable and gangue minerals; the sensitivity of valuable mineral recoveries in the rougher flotation stages to grind size, ore hardness; electricity costs; and mining vs. processing throughput constraints. Typically, copper and by-product recovery increase as grind size is reduced, which can improve revenue, but finer grinds increase power and grinding media operating costs and may have implications for tailings management. This paper explores the trade-offs that should be assessed in determining the optimum grind size, and to answer the question if coarser grinds are worthwhile. METALLURGICAL AND FLOWSHEET FACTORS Recovery-by-size behaviour of sulphide-bearing particles In base metals sulphide flotation, it is typical that flotation recoveries vary by particle size. Metal sulphide particles rendered hydrophobic by collector in intermediate size ranges (e.g. 10 to 100 um) usually show very high recoveries, often approaching 100%. Recovery usually falls away in finer size fractions (<10 um) as fine particles have little inertia and can move around bubbles without colliding and attaching to them. Coarser particles (>100 um) are susceptible to particle-bubble detachment and dropback (Jameson, 2019).

Jan 1, 2019

By C. Y. Garber

INTEREST in the sink-float plant operated by the Bunker Hill and Sullivan Mining and Concentrating Co. at Kellogg, Idaho, has recently been largely confined to the preparation and reconditioning of the galena medium. The ease and simplicity of these operations in a concentrating mill producing a flotation lead concentrate has occasioned enough comment to possibly justify a brief description. The subject is treated from an operator&apos;s viewpoint only; no effort is made to discuss it in a theoretical or technical manner. PLANT AT BUNKER HILL The sink-float plant was first operated in December 1941. It was built for the purpose of quickly increasing mill capacity with a minimum of equipment. It is the Huntington-Heberlein sink and float process using galena medium and is operated under license from the Sink and Float Corporation, New York city. Lead-zinc ore from the Bunker Hill mine is crushed to pass a I ¼ in. screen and is then washed and separated on vibrating screens having 4 ½ -mm. round-hole punched plates. The screen undersize, amounting to 33 per cent of the mill feed, is dewatered and sent to the ball-mill feed bins. The screen oversize goes to the sink-float plant, where 50 per cent of the oversize, or 33 per cent of the mill feed, is float, which is washed free of medium and disposed of as a tailing. The remaining 33 per cent is sink, which, after washing, is crushed to pass ½ in. and joins the undersize in the ball-mill feed bins. The normal plant capacity is 1700 tons per 24 hr. About 1.7 per cent of the total lead and 3.3 per cent of the total zinc in the feed is lost with the sink-float tailings (Table I). [ ] More complete descriptions of the H. H. process and Bunker Hill sink-float plant are given in several recent articles, listed at the end of the paper. PREPARATION OF MEDIUM AT BUNKER HILL New medium is made intermittently as needed, usually 12 hr. in each 6 days. A small portion of the current mill flotation concentrate is diverted from the lead cleaner cells to a I-in. vertical sand pump and pumped to a two-cell 36-in. Agitair flotation machine, where NaCn and ZnSO4 are added and the grade of concentrate is improved as shown in Table 2. Tailing from this operation is returned by gravity

Jan 1, 1946