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این مقاله انگلیسی ISI در نشریه Ceramics در 5 صفحه در سال 2010 منتشر شده و ترجمه آن 11 صفحه میباشد. کیفیت ترجمه این مقاله ویژه – طلایی ⭐️⭐️⭐️ بوده و به صورت کامل ترجمه شده است.

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چکیده

مقدمه

آزمایشات

مواد

روش ها

نتایج و بحث

نتیجه گیری

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نتیجه گیری
بدلیل مشابه بودن خواص کریستال-شیمیایی سدیم فلدسپات و پتاسیم فلدسپات استفاده از آمین و محیط HF جهت جدایش انتخابی این دو کافی نمی باشد. با این وجود استفاده از نمک ساده NaCl انتخاب پذیری بین کانی های فلدسپات را امکان پذیر می سازد.
اطلاعات مربوط به فلوتاسیون نشان می دهد که جدایش سدیم و پتاسیم فلدسپات از گرانیت هوازده با استفاده از 1 کیلوکرم بر تن HF ،در غلظت آمین 50 تا 125 گرم بر تن و استفاده از NaCl در غلظت 5 تا 15 گرم بر لیتر به بهترین شکل صورت گرفته است.در این شرایط بهترین جدایش ارتوکلاز از آلبیت بدست آمده که عیار K2O 8.43 درصد و عیار Na2O 4.19 درصد و بازیابی آنها به ترتیب 70.18 و 28.27 درصد بوده است. این کنسانتره ها برای صنایع شیشه و سرامیک مناسب است. تحقیقات در حال گسترش است تا توانایی این تکنیک را در Ph خنثی و بدون استفاده از HF شناسایی کنند.

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CONCLUSIONS

Since sodium and potassium feldspars exhibit the same crystal-chemical properties, the selective flotation between Na and K feldspars is insufficient using amine at HF medium. However, the use of a simple salt NaCl enabled the selectivity between the feldspar minerals.

Flotation data indicate that separation of Na- and K-feldspars from weathered granites with 1 kg/t HF and 50-125 g/t amine concentration is possible using NaCl in the range of 5-15 g/l. Under these conditions, the most efficient separation of orthoclase from albite is achieved with assays of 8.43 % K2O and 4.19 % Na2O and recoveries of 70.18 % and 28.27 % respectively; this concentrate is suitable for both glass and ceramic industry. Research is in progress to identify the ability of this technique at natural pH or without the use of HF.

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 چکیده
1 مقدمه
2 روش
1 2 کلیات
2 2 مدل بهینه سازی
3 2 شیوه حل مسئله
1 3 2 قاعده تصمیم
2 3 2 مکانیزم اختلال
3 پیاده سازی روش عملیات چند چاله ای
1 3 کلیات عملیات
2 3 مورد پایه
3 3 پارامترهای بهینه سازی
4 3 مورد 1 چند فرایندی چند چاله ای
5 3 مورد 2 مسئله چند فرایندی چند چاله ای با آلترناتیوهای عملیاتی در آسیاب ها
4 نتایج

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برنامه ریزان معدن، مجتمع های معادن را بادرنظرگرفتن این مسئله بهینه می نمایند که مدلهای زمین شناسی درونیابی شده، توانایی شان در زمین را به درستی نشان داده و عدم قطعیت زمین شناسی مرتبط با رسوبات معدنی را نادیده می گیرند. طبق تعریف روشهای درونیابی رسوبات و نهاشت های تیپ متوسط تولید می کنند. با توجه به این مسئله که غیر خطی بودن تابع انتقال که بین گریدهای بلوک های معدنی و خروجی های اقتصادی مجتمع معادن ارتباط برقرار می کند، استفاده از رسوبات و نهشت های تیپ متوسط در بهینه سازی، مثل شیوه های صنعتی استاندارد، راه حلی با عملکرد ضعیف روی مجموعه سناریوهای ممکن تولید می کند. در مقابل، فرمولاسیون های مبتنی بر ریسک از عدم قطعیت محلی توام یا مشترک با در نظر گرفتن شبیه سازیهای تصادفی کانسار استفاده می کنند که می تواند منجر به راه حل هایی با پاداش بالاتر و ریسک کمتر گردد.
روش مبتنی بر ریسک برای بهینه سازی مجتمع های معادن متشکل از چاله های متعدد، تلمبارها، نیازمندیهای اختلاط، مسیرهای پردازش، آلترناتیوهای عملیاتی و سیستم های حمل و نقل مطرح و پیاده شده است. به خاطر بهره گیری از عدم قطعیت زمین شناسی و سایر اجزای مجتمع معادن، راه حل های تولید شده با این روش برتر از راه حل های بدست آمده با استفاده از رویکردهای قطعی و جبری متداول از لحاظ NPV مورد انتظار بوده و اهداف تولید را تامین می نماید. راه حل های بدست آمده، توالی استخراج بلوک های معدنی در چاله های مختلف، آلترناتیو عملیاتی پیاده شده در هر مسیر پردازش و سیستم حمل و نقل بکاررفته برای حمل مواد پردازش شده به استوک های نهایی را تعریف می کنند.

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. Introduction A mining complex can be interpreted as a supply chain system where material is transformed from one activity to another (Goodfellow, 2014). The primary activities (or stages) consist of: mining the materials from one or multiple sources (deposits); blending the material considering stockpiling; processing the material in different processing paths accounting for multiple operating alternatives; and transporting the products to port or final stocks using one or multiple transportation systems. For a given processing path (e.g. mill-roaster in a refractory ore operation), it is possible to have multiple operating alternatives; for example, a mill may be operated using two different options: fine or coarse grinding (Fig. 1). If the mill is operated using fine grinding, there is often a very high energy consumption, which is associated with a higher processing cost and also requires larger residence times for the material processed, thus limiting the mill throughput. A coarse grinding option requires less energy and residence time in the mill, which decreases the operating cost and increases the mill throughput, however, it results in a lower recovery in the roaster downstream. Furthermore, different processing alternatives often impose different blending requirements. For example, the tolerable amount of free silica of the input material may be lower when operating the mill at fine grinding given that the presence of this element increases the hardness of the material. When a mill is bottlenecking the system, it is better to use a coarse grind with higher throughput in the early periods of the life-of-mine (LOM), and, to use a finer grind to maximize recovery towards the end of the LOM (Whittle, 2014). During the early periods, a mining complex incurs an opportunity cost for having material with large residence times in the mill, however, as the quantity of ore remaining in the mining complex diminishes, there is no opportunity cost. Once the material is processed through the different processing paths and using some available operating alternatives, existing transportation systems, either continuous (belt conveyors, pipe transport) or batch (trucks, rail transportation), are used to transport the processed material to one or several ports or final stocks. Accounting for transportation systems in the optimization of mining complexes is important, given that they may limit the overall system output. In a mining complex, it is common to have multimodal transportation that involves the use of separate contractors or operators for each type of transport (Zamorano, 2011). To account for the demand of transportation of material processed, it is necessary to establish the feasible relations between processing paths and transportation systems; specifically, a particular transportation system may be able to handle output material from some of the available processing paths: For example, in a pyro/hydrometallurgical complex, a hydraulic pipe may be able to transport the material output from the lixiviation plant whereas the material output from the pyrometallurgical plant is transported to the final stocks via trucks. Once the feasible transport relations are established, the demand for transportation is evaluated by considering the throughput relationships (output/input tonnages) for each processing path given the operating alternative implemented. For example, the output/input tonnage relation and the metallurgical recovery in a gold flotation plant change if the mass pull is 4 or 7 percent (Hadler, Smith, & Cilliers, 2010). When the transportation of processed material is the bottleneck in the overall system, the operating conditions at the different processing paths must be evaluated. To overcome this limitation, it may be useful to reevaluate throughput specifications of the processed material. Whittle (2010) shows that by increasing the copper concentrate from 28 to 32 percent in some periods on a sulfide deposit, the metallurgical recovery decreases by 7 percent, but the NPV increases by 6 percent given the possibility of transporting more concentrated ore on the pipe, which is the bottleneck of the system. Optimizing mining complexes by considering geological uncertainty and the different activities simultaneously is a large combinatorial optimization problem (Fig. 2).

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چکیده
1 مقدمه
2 روند آزمایشی
3 نتایج و بحث
4 نتیجه گیری

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رفتار چگال شدگی و توسعۀ میکروساختار بوکسیت هندی بسیار وابسته به ترکیب شیمیای است.هر دو نوع بوکسیت هندی مورد استفاده در این بررسی به صورت گبس در طبیعت بودند.بوکست-1 با ناخالصی کمتر همشه چگالی توده ایی و چگالی واقعی  بیشتری نسبت به بوکسیت-2دارد.هر دو بوکسیت زنتر دارای کورندوم به عنوان فاز کریستالی همراه با مقدار قابل توجه فاز ششه ایی و FeAlTiO5 با ذوب پایین هستند. تمام این فازها خواص دمای زیاد بوکست را کاهش می دهند.RUL این بوکسیت ها بین 1400-1450 درجۀ سانتی گراد بود.بوکسیت-1 مزان RUL نسبتا بیشتری نسبت به بوکسیت-2 نشان می دهد.مشاهده شد که جلوگری از فاز FeAlTiO5 باعث افزایش این RUL به 1550-1600 درجۀ سانتی گراد می شود.

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 چکیده
۱ مقدمه
۲ زمین شناسی بتول بلت
۲ ۱ تناوب هیدروترمال
۲ ۲ کانی سازی سولفیدی
۳ کانی های شاخص
۳ ۱ اسپینل روی
۳ ۲ استرولیت روی
۳ ۳ بیوتیت حاوی روی
۳ ۴ المنیت روی
۴ بحث
۵ نتیجه گیری

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کانی سازی روی- سرب و مس در  منطقه کمر بند بتول دارای منشا آتشفشانی است که در آن  توده های کانسنگ یا به صورت ورقه ای بر روی کف دریا ئ یا به صورت جایگزینی های   زیر سطح دریا  و یا شامل هر دو فرایند  قرار  دارند.  مناطق کانسنگ  مربوط به سنک های آتش فشانی فلسیک دگرگونی می باشد که دارای ورقه های مشابه سنگ شناسی است. کانی  های  دارای روی که به عنوان کانی های شاخص برای  کانی سازی سولفید فلزی پایه از کمر بند بتول پروتروزوییک  در هند مرکزی عمل می کنند،  اهمیت ژنتیکی بالایی دارند.  کانی سولفید در بسیاری از سنگ های آتشفشانی فلسیک بوده و دارای تشابهاتی با  نهشته های سولفیدی ماسیو آتشفشانی در دیگر بخش های جهان می باشند. تغییر هیدروترمال سین ولکانیک امروزه با  هسته منیزیم- کلسیم داخلی  زون بندی شده و  بخش بیرونی با   مجموعه کانی های غنی  آلومینیوم و آهن همپوشانی دارند. بیشتر  آن ها دارای کران ها و یا محدوده های چینه شناسی می باشند که  اعماق آن ها متغیر بوده و ورقه های نیمه موازی نظیر  توده های کانسنگ تشکیل شده از کانسنگ های ماسیو و نیمه ماسیو دارند وجود این کانی ها در رخنمون های سنگی مکمل داده های ژئوفیزیکی و ژئوشیمیایی  طی مرحله اکتشاف بوده و به شناسایی مناطق هدف برای حفاری کمک می کنند زیرا این کانی ها به عنوان حامل های مستقیم برای کانسار عمل می کنند.

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Introduction Indicator minerals are mineral species that, being mechanically and chemically resistant to weathering when found in clastic sediments, indicate the presence of a specific type of mineralization in bedrock (McClenaghan 2005). Their physical and chemical characteristics, including a relatively high density, facilitate their preservation and identification and allow them to be readily recovered from sample media such as till, stream sediments or soil, saprolite float and even in gossan. Indicator minerals have become an important exploration method in the past 20 years and now include suites for detecting a variety of ore deposit types including diamond, gold, Ni–Cu, PGE, porphyry Cu, massive sulphide and tungsten deposits. Investigation for base metal sulphide mineralisation, particularly for zinc has been carried out by Geological Survey of India in different prospects of the Betul Belt such as Bhawra-Tekra, BargaonTarora, Kehalpur, Chopna-Munrai, BanskhapaPipariya, Ghisi, Muariya, Koparpani, Dehalwara and Bhuyari. Presence of gahnite (ZnAl2O4) or zincian spinel in the bedrock, soil or stream sediments is considered to be an important indicator for the proximity to zinc mineralization and has been successfully used in the past to locate new areas with sulphide mineralization. Zincian spinel is reported from metamorphosed hydrothermally altered felsic volcanic rock from Muariya, Banskhapa, Koparpani and Bhuyari prospects (Mahakud et al 2001; Ghosh et al 2006; Praveen and Ghosh 2007). Recently, Ghosh and Praveen (2007) added new data from Koparpani base metal prospect of this belt. With the use of in situ microbeam techniques other minerals like zincian staurolite, Znbearing biotite and zincian ilmenite can be used as indicator minerals to locate zinc mineralization. Although minerals like staurolite, biotite and ilmenite are widespread in the alteration zones of Betul Belt, Zn-bearing staurolite, biotite and ilmenite typically occur within the alteration zones in close proximity to the sulphide-rich zones as documented by Praveen and Ghosh (2007) and Ghosh and Praveen (2007). The identification of Zn-rich minerals has major significance in exploration in Betul Belt as they occur in close spatial relationship with the sulphide rich zones and therefore act as direct vectors to ore. The present contribution was prompted by the invariable presence of one or all of these minerals within or adjacent to the sulphide rich zones in all the prospects of Betul Belt and it discusses their importance with special emphasis on their genetic significance. 2. Geology of Betul Belt The Proterozoic Betul Belt forms a conspicuous lithotectonic unit of the Central Indian Tectonic Zone (CITZ) between Mahakoshal Belt in the north and Sausar Belt in the south. Quartzite, meta-pelite, bimodal volcanic rocks (basaltrhyolite), meta-exhalites, calc silicate rocks and banded iron formation constitute the supracrustal rocks of this belt. The belt is traversed by several ENE-WSW trending ductile shear zones having sub-vertical to steep dips, which were developed during deformation (Roy and Prasad 2001, 2003). The Betul Belt represents a unique litho-package in the CITZ comprising a prominent bimodal volcanic sequence in which felsic volcanic rocks are dominant (figure 1). Sulphide mineralisation is hosted by the felsic volcanic rocks and has similarities with volcanic-hosted massive sulphide deposits (VHMS) in other parts of the world. Bimodal volcanics in the region consist of both meta-basalt and meta-rhyolite. Mafic volcanics occur as elongate lenses and often show inter- fingering relationship with the felsic volcanics. They consist of pillowed and non-pillowed metabasalts with associated mafic volcaniclastics and cherts. Felsic volcanics are represented by rhyolites, rhyolitic volcaniclastics and cherts. In mineralised areas, syngenetic hydrothermal alteration and subsequent metamorphism has obliterated primary volcanic textures in most of the rocks, however Indicator minerals for sulphide mineralisation in Betul Belt 523 relatively unaltered rocks display relict textural features based on which different volcanic facies could be recognised. The most abundant felsic volcanic facies are massive rhyolite and felsic volcaniclastics. Minor bands of laminated tuffs and chert also occur within the felsic volcanics. Massive rhyolite is generally quartz-porphyritic with rounded quartz phenocrysts in a recrystallised quartzo-feldspathic groundmass. Rhyolites at places show flow banding defined by alternate mm-cm scale pale siliceous bands and darker phyllosilicate rich bands. Volcaniclastics generally contain angular to rounded clasts in a finer grained recrystallised quartzo-feldspathic groundmass. Clasts are composed of rhyolite fragments or quartz and range in size from lapilli to bombs. Felsic volcaniclastics appear to be the favourable host rock for mineralization in most prospects. Volcaniclastic facies comprising of coarse (lapilli to bomb sized) rhyolite fragments form host rock for mineralization at Muariya prospect. Similar coarse felsic volcaniclastics form the host rocks in Bhanskapa-Pipariya and Bhanwara-Tekra prospects.

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 چکیده
۱ مقدمه
۲ آزمایش
۲ ۱ مواد
۲ ۲ روش ها
۲ ۲ ۱ فرایند فروشویی
۲ ۲ ۲ تفکیک و آماده سازی فریت های منیزیم فلزی دوپ شده
۲ ۳ تجزیه تحلیل
۳ نتایج و بحث
۳ ۱ فروشویی اسیدی اتمسفری
۳ ۱ ۱ اثر غلظت اسید اولیه
۳ ۱ ۲ اثر نسبت مایع به جامد
۳ ۱ ۳ اثر زمان فروشویی
۳ ۲ جدا سازی
۳ ۳ بازیابی فلز
۴ ۳بازیابی فلز کلی
۴ نتیجه گیری

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 فرایند هیدرومتالورژی جدید برای تولید  فریت های منیزیم دوپ شده فلزی تک فازی از ترکیبات لاتریت با استفاده از  روش های  هم رسوبی و فروشویی  اسید اتمسفری ارایه شد. غلظت اولیه HCL  برابر با 2.75 مول، برای تصفیه  سنگ معدن لاتریت با نسبت مایع به جامد 10 میلی لیتبر  بر گرم، دمای 100 درجه و زمان فروشویی 60 دقیقه  در نظر گرفته شد. تحت شرایط بهینه،  کارایی فروشی Fe, Ni, Co, Mn و Mg به ترتیب 94.5%, 96.7%, 84.5%, 85.5% و 75.2 درصد بود.  به علاوه، ضمن رسوب هم زمان Ni, Co, Mn, Mg و Fe ، بیشتر الومینیوم و  کروم در محلول فروشویی را می توان از رسوب  به دلیل رفتار های مختلف آن ها در اسیدیته 13 در محلول تفکیک شد. به علاوه  فریت های منیزیم فلز تک فازی به طور موفق از محلول های فروشویی تهیه شد به خصوص  زمانی که Msaprolite/limonite در 7:3 با غلظت اسید اولیه 2.75 و 3 مول به ترتیب کنترل شد.  تحت شرایط فروشویی بهینه، بازیابی فلزی کل حدود 89.20 درصد بود. فرایند جدید قادر به کشف مسیر استفاده کارامد و جامع   از سنگ معدن لاتریت بود.

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Introduction At present, with the reduction of the world’s nickel sulfide ore and rising demand for nickel, more and more attention has been paid on the nickeliferous laterite ore, which is an important nickel oxide ore that amounts to about 70% of global land-based nickel resources although it has lower grade and is difficult to treat (Dalvi et al., 2004; Mudd, 2010). According to the chemical analysis, the laterite ore contains many valuable metals (such as Fe, Co, Mn and Mg etc.) besides Ni, and the laterite ore can be classified into three different kinds including limonite, transition and saprolite laterite ore due to its different element composition contents. For example, the limonite laterite ore has the characteristic of high iron and low magnesium contents, while the saprolite laterite ore has low iron and high magnesium contents, which would lead to different treating methods. Generally, traditional hydrometallurgical processes of the laterite ore including pressure acid leaching and atmospheric acid leaching have attracted considerable interests because low Ni grades make pyrometallurgical processes uneconomical. For pressure acid leaching process, the most important characteristic is selective extraction of Ni and Fe (Georgiou and Papangelakis, 1998; Rubisov et al., 2000). Even so, the process has been limited in the industrial production because of the harsh leaching conditions (250–270 °C, 4–5 MPa) and numerous engineering problems. Compared with pressure acid leaching process, atmospheric acid leaching process has the advantages of high extraction efficiency, low energy consumption and low equipment cost etc. (Li et al., 2012; Luo et al., 2009; McDonald and Whittington, 2008a, 2008b). However, as for acid leaching process, it is worth noting that many impurity ions (for example, Fe3+, Mg2+ and Mn2+, etc.) besides Ni2+ are co-existing in the acid solution after leaching process, and importantly, the content of impurity ions is even higher than that of Ni2+ in the leaching solution, which might lead to nickel loss and difficult separation (Chang et al., 2010; Liu et al., 2010; Luo et al., 2010; Wang et al., 2011; Zhu et al., 2010). Therefore, how to increase nickel separation efficiency from acid leaching solution, and in the meantime make full use of valuable metals are still big challenges for effective utilization of laterite ore. Spinel ferrites such as NiFe2O4, and MgFe2O4 (formula of MFe2O4), have attracted considerable interests and efforts due to their novel magnetic and electric properties and great applications in the fields of ferrofluids (Pileni, 2001), catalysts (Lin et al., 2011), magnetic highdensity storage (Srinivasan et al., 2009), etc. Generally, ferrites or metal-doped ferrites were synthesized from pure chemical reagents with controlled Fe to M (Ni, Co, Mn and Mg etc.) mole ratio (RFe/M) of 2.0. So far, none of the effort has been taken on the synthesis of ferrites from laterite ore, especially producing metal co-doped magnesium ferrites directly from the laterite ore. Recently, by using a pressure acid leaching–hydrothermal coprecipitation–calcination method (Gao et al., 2014), we synthesized Co–Mn–Mg–Al co-doped nickel ferrites only from saprolite laterite ore. We found that the acid leaching solution with controlled mole ratio of Fe to Ni for preparation of metal-doped NiFe2O4 can be realized by selective extraction of Fe and Ni from the ore. However, this method requires two hydrothermal processes comprising leaching and hydrothermal coprecipitation, which may limit its application in the practical utilization. As mentioned above, it is well-known that relatively higher metal extraction efficiency from laterite ore can be realized by atmospheric acid leaching process, and the process is good for industrialized production. In addition, the limonite and saprolite laterite ores have the opposite characteristic of Fe and Mg contents. Considering the chemical composition contents of saprolite and limonite laterite blends (saprolite to limonite mass ratio 1:1), it is reasonable to assume that the leaching solutions with controllable RFe/M of about 2.0 can be obtained by atmospheric acid leaching of laterite blends and M is the sum of the molar concentrations of the metal ions. In this paper, synthesis of metal co-doped magnesium ferrites from laterite blends was investigated by using an atmospheric acid leaching– coprecipitation method. During the leaching process, the effects of the critical factors including the initial acid concentration, liquid to solid ratio and leaching time on metal leaching efficiencies were studied. Then coprecipitation method was adopted to separate Ni, Co, Fe, Mn and Mg from Al and Cr in the leaching solution by adjusting the pH value due to their different precipitate behaviors.

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چکیده
۱مقدمه
۲ مواد و روش ها
۲ ۱ نمونه ی ذغال سنگ و شرایط کشت سویه های باکتریایی
۲ ۲ ازمایش گوگرد زدایی زیستی ذغال سنگ
۲ ۳ ازمایش فرو شویی زیستی پیریت
۲ ۴ تجزیه تحلیل نمونه
۳ نتایج و بحث
۳ ۱ فرایند گوگرد زدایی زیستی زغال سنگ
۳ ۲عملکرد فروشویی زیستی پیریت A caldus
۳ ۳ تجزیه تحلیل پیریت توسط میکروسکوپ الکترونی نگاره
۳ ۴ خصوصیات سطحی سلول ها و پیریت
۳ ۵ تجزیه تحلیل فروشویی زیستی پیریت و گونه زایی گوگرد
۴ نتیجه گیری

بخشی از ترجمه:

 سویه اسید دوست و گرما دوست A. caldus نخست برای گوگرد زدایی زیستی زغال سنگ استفاده شد . نتایج نهایی حاکی از گوگرد زدایی زیستی پیریت و گوکرد زدایی زیستی کل به ترتیب 47 و 19 درصد بود. بعد از فراوری به مدت 40 روز، سلول به سطح پیریت هاچسبیده و هیچ گونه خوردگی روی سطح کانی وجود نداشت. به علاوه، سلول های رشد یافته سوبستراهای مختلف گوگرد، خصوصیات سطحی متفاوتی رانشان داد ند. بر عکس، سلول های رشد یافته روی پیریت و گوگرد عنصری دارای گروه های عاملی اب دوست تر از تیوسولفات بودند.
گوگرد عنصری یک عنصر و مولفه مهم گوگردی در فروشویی زیستی زغال سنگ می باشد با این حال هیچ گونه ترکیب معدنی ثانویه شناسایی نشد.

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Introduction Coal is the most abundant fossil fuel in the world. However, the combustion of coal containing high sulfur causes serious environmental damage such as acid rain because of emission of sulfur dioxide [1]. Sulfur is present in coal mainly in three forms: pyritic, organic and sulfate sulfur. The organic sulfur is the integral part of the coal matrix, whereas pyritic sulfur is present in coal as mineral matter. There are several physical–chemical methods employed to eliminate the organic or inorganic sulfur fractions of coal by floatation, oxidation and reduction with chemicals [2]. Nevertheless, these physical methods cannot remove the pyrite when these sulfides are finely dispersed in coal matrix. Furthermore, although the chemical methods can remove organic and inorganic sulfur from coal, which will also produce hazardous secondary products and lead to loss of partial combustible matter [3,4]. Biodesulfurization offers a clean alternative method to remove sulfur from coals, in which the microbes can catalyze the biochemical reaction in an aqueous medium resulting in the oxidation and dissolving of the sulfur content into sulfate [4,5]. Among of them, the mesophillic microbes, mainly A. ferrooxidans and A. thiooxidans have been frequently applied to remove the pyrite from coals [4,6,7]. However, in the last decade, limited literatures reported that the thermophilic microbes were used to eliminate the sulfur from coal [8,9]. Recently, A. caldus has been reported as the dominant sulfur-oxidizing thermophilic bacterium accounting for the relatively high percentage in bioleaching pyrite at 45 °C [10,11]. To our knowledge, it has never been used in coal’s biodesulfurization process. It is well known that the attachment of cells on the surface of mineral is reported to be involved in the oxidation of sulfide minerals, which depends not only on the interfacial process between the cells and mineral surface but also the biochemical properties of the cells. Furthermore, it has been reported that the surface properties of bacterial cells are significantly influenced by the growth conditions [12–14]. Hence, a better understanding of the surface properties of functional bacteria under different conditions is helpful to illustrate the interfacial interaction between cells and mineral surface [15,16]. The bio-oxidation of pyrite has been widely investigated and a possible mechanism has been proposed: the stepwise oxidation and the formation of surface-bound thiosulfate, and then oxidized via tetrathionate and other polythionates, finally to sulfate [17–19]. Many investigators agreed that the precipitation of mineral phases during the pyrite oxidation process may affect the efficiency of sulfur removal [20,21]. Besides the reported jarosite, the sulfur-rich layers may also deposit onto the surface of pyrite, thus preventing the contact of the microorganisms and hindering the leaching process [21–23]. The object of this study is to investigate the coal’s biodesulfurization and the coal pyrite’s bioleaching by thermophilies A. caldus cells and the interfacial interaction between cells and coal pyrite surface with the approaches of zeta-potential, FT-IR spectroscopy, SEM and sulfur Kedge XANES. 2. Material and method 2.1. Coal sample and bacterial strains’ culture conditions The coal sample used in the experiment was collected from Indonesia area named MTWL-7 and the total sulfur of it was 2.15%, which was sterilized at 120 °C for 20 min. The pyrite powders used in the experiments were picked up from the coal samples and the component was analyzed with XRD (data not shown). A. caldus derived from the acid hot spring of Yun-nan province, south east of China was cultivated in the Starkey basal salt medium [24] supplemented with sulfur powder, pyrite, coal and thiosualfate, and the initial pH of culture were 2.5, 2.5, 2.5 and 4.8, respectively. The culture of A. caldus was grown in 250 mL flasks containing 100 mL medium and incubated at 40 °C with 170 rpm shaking. Bacterial cells were cultured until late exponential phase and harvested by the following procedure: the cultures were filtered three times through filter paper (pore size 7–10 μm) to remove the suspended solid particles and then the filtrates were centrifuged at 10,000 rpm for 15 min, and the cell pellets were washed thrice in diluted sulfuric acid (0.1 mol.L−1 ) in order to remove any trapped ions, finally all the samples were freezedried and stored at −20 °C for zeta-potentials and the FT-IR test. 2.2. Coal biodesulfurization experiment The biodesulfurization processes were carried out in flasks with a volume capacity of 100 mL in 250 mL of basic salt medium, 10% w/v pulp density, particle size of −65 Tyler mesh, the initial cell concentration was 1.0 × 106 cells.mL−1 , processing time of 30 d. Every four days, the pH values, and concentration of iron of biodesulfurization system were determined. At the end of the experiments, the coal samples were prepared with the method mentioned in the literature [25], then the total sulfur and the relative sulfur contents in the coal samples were measured [26].

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فهرست مطالب:

روش های آزمایشی
مواد
مواد و روش ها
نتایج و بحث
اثر تیتانیوم اکسید بر ویسکوزیته
نتیجه گیری

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در این مطالعه، اثر تیتانیوم اکسید بر ویسکوزیته ناحیه تماما مایع سیستم CaO–SiO2–17mass% Al2O3–10mass% MgO–TiO2 در نسبت های مختلف CaO/SiO2، همبستگی مستقیم و مثبتی با ساختمان پسماند فلز حرارتی داشت.
افزایش تیتانیوم اکسید منجر به کاهش ویسکوزیته با تاثیر بر ساختار شبکه سیلیکات شد که ورقه های پیچیده سیلیکاته به ساختار های ساده تر شکسته تر شدند با این حال ساختارهای آلومینات به نظر می رسد متاثر از افزایش تیتانیوم اکسید نمی باشند. هر دو آنالیز FTIR و رامان حاکی از دپلی مریزاسیون پسماند بوده و مقادیر ویسکوزیته را به مقدار تیتانیوم اکسید به طور مستقیم ارتباط می دهند. به علاوه، افزایش نسبت CaO/SiO2 به نظر می رسد در کاهش ویسکوزیته در مقایسه با افزایش تیتانیوم اکسید کارامد تر باشد. با توجه به وابستگی درجه حرارت نسبت 1.2 CaO/SiO2، انرژی اکتیواسیون جریان ویسکوز بین 226 و 260 کیلوژول بر مول بسته به ترکیب فلز می باشد.

بخشی از مقاله انگلیسی:

 Introduction Cost issues related to integrated steel mill operations and unavailability of high grade raw materials with high concentrations of iron has sparked interest in the use of other potential sources of iron bearing materials once regarded as uneconomical for the present day integrated steel mill operations. In particular, iron ores consisting of significant TiO2 such as ilmenite is comparatively abundant in Australia and other parts of the world. These ores due to the high TiO2 content may result in significant issues such as lower reduction degree of the ore,[1–4] variations in the liquidus temperature of the slag,[5–7] reduced refining capacity of the slag,[8–10] and slag viscosity deviations from optimum processing conditions.[11–15] However, potential benefits of TiO2 in the blast furnace slags have also been realized from work done by Morizane et al.[9] and others.[15–18] By forming a protective layer of titanium carbo-nitride on the refractory brick, the premature failure and erosion of the hearth was found to be inhibited. Thus, both the thermophysical and chemical properties of the slag can change with TiO2 additions and a fundamental understanding on these topics, in particular the effect of viscosity, is essential to maintain and optimize the overall efficiency of the blast furnace operations. Handfield and Charette[12] showed small TiO2 additions can decrease the slag viscosity, but large TiO2 additions had the opposite effect. Ohno and Ross[13] described TiO2 additions to increase slag viscosity in the CaO–SiO2–Al2O3– TiO2 slags under reducing conditions. Work done by Shankar et al.[19] showed the effect of TiO2 in the CaO– SiO2–MgO–Al2O3 slag system under Ar, where TiO2 up to 2 mass% decreased the viscosity at various fixed extended Vee ratio [(CaO þ MgO)/(SiO2 þ Al2O3)] between 0.46 and 0.83. Saito et al.[14] also studied the effect of TiO2 in the CaO–SiO2–MgO–Al2O3 slag system at 10 and 20 mass% TiO2 content. TiO2 lowered the viscosity of the slag and the corresponding activation energy of viscous flow decreased with TiO2 content. Thus, slag viscosity can be affected by the amount of TiO2 additions. However, to the knowledge of the authors, the viscosity data between 0 and 10 mass% TiO2 is not well defined for blast furnace type slags and its correlation with the slag structure has yet to be fully understood. Therefore, understanding and controlling the effects of TiO2 on the viscosity of blast furnace type slags, when raw materials’ containing significant amounts of TiO2 such as ilmenite is utilized, would be essential in maintaining stable blast furnace operations. In this work, the influence of TiO2 on the viscous behavior of calcium-silicate based slags with 17 mass% Al2O3 and 10 mass% MgO based slags have been studied using a rotating viscometer. To directly correlate the relationship between slag structure and viscosity, flux samples from the fully liquid region at 1773 K were quenched and analyzed using Fourier transform infra-red (FTIR) and Raman spectroscopy.2. Experimental Methods 2.1. Materials Slags were prepared using reagent grade chemicals of CaO, SiO2, MgO, Al2O3, and TiO2. Using a Pt-10 mass% Rh crucible, slags were premelted under 0.4 L min1 of Ar to obtain a homogeneous mixture of CaO–SiO2–17 mass% Al2O3–10 mass% MgO. After 3 h, the homogenized samples were quenched and crushed for the primary experiments. Chemical composition of the post-experimental slags was analyzed using X-ray fluorescence (XRF) spectroscopy (S4 Explorer; Bruker AXS GmbH, Karlsruhe, Germany) and no apparent change was observed. Thus, the initial weighed mixture was taken as the slag composition. The experimental chemical composition is given in Table 1. 2.2. Apparatus and Procedure The viscosity was measured using a rotating spindle connected to a calibrated Brookfield digital rheometer (LVDVIIþ; Brookfield Engineering Laboratories, Middleboro, MA). The details of the experimental apparatus have been given elsewhere,[20,21] but a schematic diagram is shown in Figure 1. Samples weighing 120 g were prepared according to Table 1 and placed in a Pt-10 mass% Rh crucible (H: 60 mm, ID: 40 mm). Furnace temperature was calibrated using a reference B-type thermocouple and controlled within 3 K using a PID controller. Slag samples were heated to 1773 K and sufficiently held for more than an hour in an Ar gas atmosphere to achieve thermal equilibrium. Measurements were taken at each target temperature during a 58C min1 cooling cycle after 30 min of soaking and in the fully liquid region of the slag system well above the break temperature (TBr) of each slag composition. The break temperature has been described in detail by Kim and Sohn[21] and Sridhar et al.[22] Viscosity measurements taken at temperatures higher than the break temperature are assumed to be in the fully liquid region of the slag system, of which the present experimental measurements have been done. The break temperature can easily be identified using the example given in Figure 2 for the CaO–SiO2–17 mass% Al2O3–10 mass% MgO–5 mass% TiO2 slag at constant CaO/SiO2 of 1.2. The natural logarithm of viscosity (ln h) as a function of reciprocal temperature (1/T) shows a significant increase below temperatures of 1598 K, which is the starting temperature for the formation of large solid precipitates. Viscosity measurements were also taken during a heating cycle of 58C min1 and 30 min of soaking time, but no appreciable difference was found during measurements in the fully liquid region of the slag system and thus subsequent viscosity measurements were done during the cooling cycle.

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فهرست مطالب:

چکیده
۱ مقدمه
۲ آزمایش
۱ ۲ مواد
۲ ۲ روش ها
۳ نتایج و بحث
۱ ۳ تاثیر مواد کف زا
۴ نتیجه گیری

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 نتایج حاصله در این مقاله، به شرح ذیل است:
تغییر حالت با شدت بالا به عنوان مرحله پیش تیمار موجب بهبود ریکاوری فلوتاسیون ذرات طلا تا 24 درصد، افزایش سرعت آن تا 2 تا 3 برابر و گرید 50 درصد خواهد شد.
نتایج حاصله را می توان با افزایش در غلظت ذرات شناور در نتیجه تجمع ذرات ریز طلا بر اساس فلوتاسیون حامل توجیه کرد. این را می توان با افزایش در سرعت فلوتاسیون حقیقی، منحنی های گرید-ریکاوری و سرعت ترکیب پایین نشان داد .
لزوم استفاده از حالت دهنده های جدید یک عامل کلیدی برای بهبود ذرات ریز و ریکاوری آن ها با فلوتاسیون کف می باشد و حالت دهنده های برشی بایستی اختراع شوند.

بخشی از مقاله انگلیسی:

been reported on in the past (Trahar and Warren, 1976; Fuerstenau et al., 1978, 1988; Kitchener, 1978; Fuerstenau, 1980; Trahar, 1981; Sivamohan. 1990) focusing on the problems posed by very fine particles during flotation, and several alternative processes have been suggested. Most of them are based on particle aggregation, namely shear flocculation, carrier/autogenous carrier flotation, oil agglomeration and selective flocculation (Rubio and Kitchener, 1977; Kitchener, 1978; Guerra et al., 1986; Rubio and Marabini, 1987; Subrahmanyan and Forssberg, 1989; Warren, 199 1; Behl and Moudgil, 1993). The shear flocculation process is based on the selective aggregation of hydrophobic particles in systems under high turbulence. Warren has reported several studies on shear flocculation (Warren, 1975a,b, 1991; Koh and Warren, 1979a,b) and concluded that formation of aggregates takes place within a short agitation range. Thus, at high stirring speed, the formed aggregates redisperse, due to attrition and/or high shear stress. Autogenous carrier flotation (Hu et al., 1988) employs the same mineral particles which are being floated (but coarser) as carrier, and it is believed that this process proceeds like the shear flocculation with the fine particles floating attached to the larger particles (Chia and Somasundaram, 1983). Also, studies on ‘induced’ carrier flotation were carried out by Rubio and Hoberg (1993) to recover fine particles using hydrophobic polymeric spheres as carrier particles. Dianzuo et al. (1988) refer to shear flocculation, carrier flotation and emulsion flotation as main examples of hydrophobic aggregation enhancing particle capture by bubbles. The concept of shear flocculation has been extended to the conditioning stage ahead of flotation (Rubio, 1978; Bulatovic and Salter, 1989; Stassen, 1990, 199 1; Rubio and de Brum, 1994). These authors claim that the energy transferred in the conditioning stage, often expressed as conditioning time at constant impeller speed or as impeller speed at constant time, has a pronounced effect on the concentrate recovery, grade and flotation rate. The high intensity conditioning process, HIC, as it has been named, enhanced the flotation recovery of the very fine particles of copper sulfides (Bulatovic and Salter, 1989), gold, uranium oxide and pyrite fines (Stassen, 1990, 1991) and oxidized copper ores and copper and molybdenum sulfides (Rubio, 1978; Rubio and de Brum, 1994). The fact that carrier or autogenous carrier flotation may be the responsible or be operating in the overall phenomena have not being recognized. Because of that and because mechanisms involved are not very clear and data reported on the effect of HIC are still insufficient, the aim of this work is to report results on the characterization of the effect of HIC on flotation of gold fines. The effect of HIC was monitored by measurements of the ‘true’ flotation of gold fine particles, concentrate grade and recovery, degree of entrainment of values and flotation kinetics. The study has been conducted with gold fines from North Chile (Atacama), where gold is usually found in a finely dispersed form and, thus, fine grind is required in order to achieve particle liberation to enhance flotation recovery. This results in significant losses in the flotation of the ultrafine particle fractions. L. Valderrama, J. Rubio/Int. J. Miner. Process. 52 (1998) 273-285 275 2. Experiimental 2.1. Materials A representative ore sample of a copper/gold ore from Atacama (46 kg) was crushed to less than 10 mesh. Grinding was carried out in a stainless-steel ball mill (17.2 cm diameter and 22 cm length) with 55% solids by weight during 21.5 min to yield a material which was almost 65% under 400 mesh fraction (90% < 200 mesh). The mineralogical analysis showed that the ore consisted mainly of quartz, limonite, hematite, pyrite and chalcopyrite, and that gold liberation in the 200 mesh fraction was about 82%. The sample analyzed 3.7 g/t Au, 4.5 g/t Ag, 1.36% Cu, 14.15% Fe and 59.24% SiO,. Table 1 and Fig. 1 show the gold and particle size distribution in feed. Results show that about 64% of total gold is found below 38 km. 2.2. Methods Conditioning of a gold flotation feed (22% w/w>, before HIC, was first conducted in a 3-l Denver cell, during 3 min to allow collector and frother adsorption. Then the same cell was endowed with four acrylic baffles to produce the HIC under a turbulent regime. Results, expressed in kW h/m3 pulp, were obtained by stirring at a constant 1500 rpm, the pulp volume at 2 1 with conditioning time varying in the range of 25- 100 s to yield 0.5 to 4 kW h/m3 pulp inputs. The energy transferred was monitored using a voltmeter and an amperimeter. The HIC was performed in the presence of 120 g/t amylxanthate as collector and the frothers (72 g/t> studied were: pine oil, MIBC and DF-250. The medium pH varied between pH 7.5 and 7.8.

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چکیده
مقدمه
مطالعات فلوتاسیون در مقیاس آزمایشی
روش آزمون فلوتاسیون در مقیاس آزمایشی
کانی ها
تست فلوتاسیون
تست های فلوتاسیون برای بالک رافینگ ۱
جزییات تست ها
نتایج و بحث
نوع و دوز جمع اوری کننده
اثرات ماده پراکنده کننده
اثرات سولفیدایزر(Na2S)
اثرات روغن کاج(کف کننده)
تاثیر چگالی خمیر
اثرات زمان تغییر حالت
تسست های فلوتاسیون برای بالک رافینگ ۲
جزییات تست ها
نتایج و بحث
تست های فلوتاسیون برای پاکسازی ۱
جزییات تست ها
جمع بندی و نتیجه گیری

بخشی از ترجمه:

پارامترهای فرایند فلوتاسیون از نظر تمرکز مقدار مس یعنی کالکوپیریت کانسنگ مس وزیرستان شمال در مقیاس آزمایش برای بدست اوردن غلظت مس کافی برای فراوری های متالورژیکی بیشتر مورد مطالعه قرار گرفت. پارامترهای مهم فلوتاسیون یعنی نوع و دوز جمع اوری کننده، دوز منقبض کننده، منتشر کننده، کف ساز و زمان شرطی مورد آزمایش قرار گرفت. طی بهینه سازی گام به گام پارامترهای فلوتاسیون، مقدار مس از ۰.۹ تا ۲۰ درصد در مرحله مخلوط سازی و تا بیش از ۲۲ درصد در مرحله پاکسازی تک مرحله ای با بازیابی بیش از ۸۳ درصد روزآمد شد. فلوشیت حاوی محصولات مختلف فلوتاسون برای یک متمرکز کننده صنعتی نیز پیشنهاد شده است.

۱ مقدمه

در وزیرستان شمال، در مناطقی مثل شینکای و دگان، حجم عظیمی از نهشته ها و رسوبات مس موجود می باشد. منابع کانسنک مادون قرمز حدود ۱۲۲.۷۱ میلیون تن است که دارای محتوی مس متغیر از ۰.۸۳۶۵% تا بیش از ۲% می باشد(FATA DC 1985)(ونگ زیتان ۱۹۹۶). هدف این مقاله، مطالعه سطح و غلظت مس کانسنگ وزیرستان شمالی با محتوی مس حدود ۰.۹ درصد تا یک سطح قابل قبول برای فراوری متالورژیکی بود. بخش غربی وزیرستان دارای یک کمربند آذرین پیچیده است. این کمربند دارای شمال شرقی جنوب غربی بوده و متشکل از توده های فرا مافیک حاوی هارزبورگیت، پیروکسنیت،پریودودایت و دیونیت می باشد. انواع بیرونی یا نفوذی شامل دیوریت، کوارتز-دیوریت، دیوریت میکروکوارتز، دولریت و گابرو می باشد.

بخشی از مقاله انگلیسی:

ABSTRACT Flotation process parameters were studied to concentrate the copper content i.e. Chalcopyrite of North Waziristan copper-ore on Pilot-scale to obtain a copper concentrate suitable for further metallurgical treatment. The important flotation parameters, e.g. type and dosage of collector, dosage of depressent, dispersent and frother and conditioning time for collector were examined. During stepwise optimization of flotation parameters, the copper content was upgraded from 0.9% to 20% in roughing stage, and to as high as 22% in a single-stage cleaning, with recoveries of over 83%. A flowsheet depicting different products of flotation, for an industrial concentrator, has also been suggested. INTRODUCTION In North-Waziristan, at places like Shinkai and Degan, huge deposits of copper are available. The inferred ore-reservers are 122.71 million tons,having a copper content varying from 0.3865% to as high as 2% (FATA DC, 1985); (Wang Zhitian, 1996). The purpose of this paper is to study concentrateion of the North Waziristan copper ore (Shinkai area), having copper content 0.9% to a level suitable for metallurgical treatment. The western part of Waziristan constitutes a complex igneous belt. The belt extends NortheastSouthwest and consist of ultramafic masses, consisting of harzburgite, pyroxenites, periododites and dunites, generally serpentinized. The intrusives comprise diorites, quartz-diorites, micro-quartz diorites, granodiorites, dolerites and gabbros. Volcanics include fine-grained porphyritic pillow basalts and andesites, with subordinate breccias and minor dacite, rhyodacites, tuffs and agglomerates (Badshah 1983). To reconfirm the previously carried out studies (Khan, 1994), laboratory-scale petrographic studies of the ore showed that the sulphide mineralization is present in the form of the stock work-internally brecciated. The suphide minerals are disseminated and are in skeletal form. The pillow breccia is chloritized and contains minerals like hornblende and pyroxene. The ore-minerals are mostly cube-shaped, sporadically distributed throughout the rock but also found in the form of veinlets. Chalcopyrite exists as a major mineral, with malachite, azurite, tennantite as minor valuable minerals. Pyrite, pitchlimonite, bronchantite, graphite, chamosite, nimite and quartz exist as gangue minerals (Rafiq Mazhar, 1999). PILOT-PLANT FLOTATION STUDIES The aim of the present research is to develop an economically viable commerical process to utilize the 122 million tons of copper ore of the North Waziristan area.Pilot-scale flotation studies were carried out to upgrade its copper-content by physical concentration method to a level which is suitable for metallurgical treatment to produce blister copper. Flotation-process parameters, such as collector type and dosage, dispersent, pulp density, etc, were investigated. FLOTATION TEST-PROCEDURE AT PILOT SCALE Minerals The copper ore was from the Shinkai area (North Waziristan) and assayed at 0.9% Cu, 0.0134% Zn, 0.0126% Pb, 0.0110% Co, 0.0067% Ni, 1.740 ppm Ag, and 0.22 ppm Au. Quartz was the main gangue mineral. Communition For each test-run of the flotation-process, approximately 100 kgs of copper ore was reduced in size to 80% passing 4mm. The crushed material was wet-ground for 20 min in a closed-circuit ballmill classifier, at a solid-liquid ratio of 1:2.3 (i.e. 30% solids by weight), to 80% passing 75mm. The coarse particles were difficult to float, so the grind was chosen to be fine i.e. 75mm (Rafiq Mazhar, 1999). The fine-size grinding also promotes the self-induced flotability of the particles (Senior G.D. and Trahar W.J., 1976). Flotation Testing The product of the ball-mill was conditioned and subjected to flotation in a bank of 25 and 28 litres flotation-cells. All the reagents were added in the conditioning stage, except caustic soda (NaOH) which was added during the grinding stage as pH regulator. The conditioned pulp was subject to two roughingstage flotations and one cleaning-stage flotation. The concentrates from the Bulk Roughing-I and Bulk Roughing-II were combined and cleaned, to obtain the final concentrate. For Bulk Roughing-I, the parameters investigated were type and dosage of collector, depressant dosage, sulphidizer dosage, pine-oil dosage, dispersent dosage and conditioning time. In any stage of flotation . i.e. roughing or cleaning, while studying the effect of variation of any parameter on the grade and recovery of copper, the rest of the parameters were kept constant. (Table 1). Normal tap-water was used for conducting all the tests. All the reagents were of technical grade, except sodium cyanide (NaCN), caustic soda (NaOH) and sodium sulphide (Na2S) which were of commercial grade. Chemical analyses were carried out with X-Ray Spectrometer System (Jeol, Model JSX-603). FLOTATION TESTS FOR BULK ROUGHING-I In the flotation method of concentrating minerals, fast floating minerals are recovered in bulkroughing stages, in order to obtain maximum recovery of valuable minerals (Wills B.A, 1985). The parameters investigated during this stage were type and dosage of collector, pH value, depressant dosage, sulphidizer dosage, pulp density, and pine-oil dosage. Generally, higher dosages of reagents are used during this stage of flotation. Details of Tests Fifteen tests were carried out to evaluate the flotation- response , using different dosages and. type of collector. Five tests were conducted each to investigate the effect of varying pH and depressent dosage on the grade and recovery of copper. The effects of sulfidizer was investigated by using dosage in the range of 10- 60 gms/ton. Three tests were conducted to investigate the effect of pine oil (frother) by varying its dosage between 25 and 70 gms/ton. The effect of variation in pulp-density was investigated by varying it between 15 and 35%. Similarly, a few tests were conducted to investigate the effect of conditioning–time on the grade and recovery of copper. All the products, including tailings, recovered during. flotation test work, were dried and analyzed for copper. From the results, grade and recovery curves were drawn. The concentrate from the Bulk Roughing-I was then routed to the cleaning stage. DISCUSSION OF RESULTS Collector Type and Dosage The curves in Figure 1 (Tests 1-5) show that, with an increase in NaPX (Sodium Propyl Xanthate) there is a corresponding increase in the grade/ recovery of copper (Ralston V.E., 1991, Senior, 1991, Keble 1982, Woods, 1971, Ross, 1981, Swat 1991). Beyond the dosage of 200 gms/ ton, there is a decrease in grade and a slight decrease in recovery. This may be due to the non-specific adsorption of the collector by the gangue particles (Woods, 1971; Ross,1991) and, possibly, due to the development of collector multilayers on the particles, reducing the proportion of hydrocarbon radicals oriented into the bulk solution (Smart 1991; Fuerstenau and Palmer 1976; Smith and Akhtar, 1976), thus reducing the grade of the concentrate; the critical dosage of NaPX observed at this stage was 200 gms/ton. The curves in Figure 2 (Tests 6-10) show a similar behaviour of NaEX (Sodium Ethyl Xanthate) to NaPX. By using NaEX, the grade/recovery is almost the same upto 150 gms/ton but, later, the recovery has not improved as in case of NaPX. This may be due to the fact that hydrophobic action of Xanthate collectors decreases with the, decrease in molecular weight of the alkyl group (Wills,1985); (Majima and Takeda, 1968); (Kelly and Spottiswood, 1982). The curves in Figure 3 (Tests 11-15) show the effects of KPX (Potassium Propyl Xanthate) on the grade and recovery of copper. The curves reveal a similar behaviour of KPX to that of NaPX. By using collector KPX, there is not much effect on the grade/recovery of copper in the concentrate. Hence, Sodium or Potassium cation in collector has no effect on flotation (Crozier, 1991).

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