Altın, gümüş içeren sülfürlü, bakırlı-pirit cevherlerinden tiyoüre liçi ile altın ve gümüş kazanımı

Abstract

ÖZET Bu çalışma kapsamında, altın içeren Küre sülfurlü bakırlı-pirit cevherlerinden tiyoüre liçi ile altın ve gümüşün kazanılma olanakları araştırılmıştır. Siyanüre alternatif tiyoürenin liç kimyası, kinetiği ve uygulama çalışmaları incelenerek, sülfurlü bakırlı-pirit cevherlerine uygulanabilirliği, laboratuvar ölçeğinde araştırılmıştır. Ayrıca tiyoüre liçinden elde edilen sonuçlar, uygulamalarda kullanılan değerlerle yapılan siyanür liçi sonuçlan ile karşdaştirılmıştır. Küre sülfurlü bakırlı-piritli cevherleri üzerinde tiyoüre ile yapılan ilk araştırma niteliğindeki bu çalışmada, kullanılan cevher, 19.47 g/t Au ve 15.5 g/t Ag ve yüksek silis oranıyla, bölgede ender olarak karşılaşılabilecek kuvarsça zengin masif sülfurlü cevherleşme zonunu temsil etmektedir. Altın ve gümüşün kazanılmasına yönelik olarak, cevherde çözünen bakırın ayrılmasından sonra ve doğrudan tiyoüre liçi yapılmıştır. Liç parametrelerinin çözünme verimlerine etkilerinin araştırıldığı deneylerde, cevherde bulunan bakırın % 5-8 arasında çözündüğü, çözeltideki bakır iyonlarının tiyoüre tüketimini etkilediği belirlenmiştir. Çözünen bakırın ayrılması amacıyla yapılan seyrettik sülfürik asit-demir sülfat ile çözündürmede, asit konsantrasyonunun bakır çözünmesi üzerinde fazla etkili olmadığı, buna karşılık demir sülfat konsantrasyonunun artmasıyla çözünürlüğün azaldığı saptanmıştır. Çözünen bakırın ayrılmasından sonra yapılan tiyoüre liçi sırasında, çözeltiye geçen bakır oranının % 0.5-1.9 arasında olduğu bu grup deneylerde, altın çözünürlüğü % 40-42 olarak tesbit edilmiştir. En uygun tiyoüre liçi koşullarının saptanması için reaktif konsantrasyonu, çözelti Eh-pH değerleri, redoks potansiyelinin kontrolü ve liç süresi gibi parametreler araştırılmıştır. % 69.1 maksimum altın çözünürlüğüne; 10 g/l tiyoüre, 0.05 M demir sülfat, 0.05 M sodyum sülfit konsantrasyonlarında, başlangıç pH=l ve Eh =280 mV civarında, 4 saatlik liç süresi sonunda ulaşıldığı saptanmıştır. En iyi altın çözünürlüğü için tiyoüre tüketimi 19.6 kg/t olarak hesaplanmıştır. Cevhere doğrudan tiyoüre liçi uygulamasında, % 58.4 maksimum altın çözünürlüğü için tiyoüre konsantrasyonu 15 g/l olarak bulunmuştur. Altın çözünürlüğünün % 10 civarında düştüğü bu grup deneylerde tiyoüre tüketimi 34.4 kg/t mertebelerine çıkmaktadır. Elde edilen sonuçlara göre, bakır iyonlarının tiyoüre tüketimini arttırdığı görülmektedir. Doğrudan liç uygulamasında, kademeli reaktif ilavesiyle maksimum altın çözünürlüğü % 83'e çıkmakta, ancak tiyoüre tüketimi de artarak 55 kg/t seviyesine yükselmektedir. Liç çözeltisinde kalan tiyoürenin yeniden kullanılması halinde % 73 altın kazanma verimi gerçekleşmekte, ancak iki kademede kullanılabilen liç reaktifi, ikinci kademe sonunda tümüyle tüketilmiş olmaktadır.

Gold is the first metal to attract the attention of man for aesthetic as well as commercial purpose, and it has continued to be a commodity of considerable demand in today, because of its unique physical and chemical properties. The first working of gold data back to the late stone age; all the earliest civilizations were familiar with gold. It is likely that gold wetting by mercury (amalgamation) was known in 1000 BC, although it was not commonly used as a commercial gold recovery process until much later. In Turkey around 700 BC, the first gold coins were produced but it was only after a process for fire refining of gold was developed in 560 BC that pure gold coins were minted. Alloying of gold and silver was known to the Egyptians as early as 500 BC. Under the historical development; many countries and civilizations were met with unchangable values of precious metals, predominantly with gold, silver. Although the early treatment of gold ores were almost totally involved with gravity separation, physicochemical and chemical processes were more pronounced and several techniques also were established by the time. In 1867 Rae (USA) patented a process for cyanide leaching of gold and silver ores although this was never used. The cyanidation process as it is now known, was patented between 1887 and 1888 by McArthur and the Forrest brothers, and was rapidly developed into a commercial process, first at Crown Mine (New Zeland) in 1889. The technology spread rapidly and was used at Robinson Deep (S. Africa) in 1890, Consolidated Mercur (Utah,USA) and Calumet (California,USA) in 1891, El Ore (Mexico) in 1900, and La Belliene (France) in 1904. The characteristics of an ore deposit and its mineral assemblages determine the mining methods, extraction process requirements and, in particular, the performance of all chemical process involved in precious metals extraction.For this reason, the mineralogical properties and concentration characteristics of gold ores and also treatment products of most importance to the process engineers and hydrometallurgists, are as fallows; -Gold ore grade, and composition, -Liberation characteristics of all valuable minerals, -Gold mineral type, and grain size distribution, -Concentrations of any other valuable minerals. Native gold, gold tellurides and gold-bearing sulfides are the major economic sources of gold. Among the sulfides, pyrite and arsenopyrite are the most important, followed by chalcopyrite, galena and sphalerite. Native gold is also found in association with pyrrhotite, quartz, carbonates, carbonaceous matter and many other minerals. Several gold ore deposits, containing the above minerals individually or combined some of them, are classified as refractory when a significant portion of the gold cannot be extracted efficiently. According to the an analysis of the extraction methods used for primary gold recovery in 1989, it is indicated that more than 94 % of gold was recovered with leaching as well as gravity concentration or an intermediate flotation step. Where the gold and silver bearing ores amenable to the direct cyanidation, they are called free-milling ore. As world reserves of higher grade and free- milling precious metal ores are becoming depleted, efforts are being made to develop new technologies. The other 6 % of production came from plants that used a pre-treatment step before the application of classical gold recovery techniques because of poor gold extraction and high reagent consumption. Although the ratio of refractory ores does not seem enough high, it is gaining more important attention in gold metallurgy. The need to treat increasingly low grade and/or refractory gold ores and the continuing search for improvements in the economics of existing operations has led to several developments and innovations in gold extraction metallurgy during the last two decades. vnProcesses for the extraction of Au and Ag in terms of hydrometallurgical base have improved dramatically over the years. Cyanide the oldest one, has been recognized for long time as a powerful lixiviants for Au and Ag, forming very stable cyano complexes with both metals. Although the process is so simple, efficient and inexpensive, in fact, cyanide solutions are toxic and must be handled carefully to avoid damaging the environment. In cyanidation, cyanide solution is not efficient in leaching certain classes of gold ores which are considered refractory to cyanidation. Because of this limitation and increasingly stringent pollution control regulations, use of lixiviants other than cyanide for the recovery of Au and Ag has received considerable attention during the past two decades. Among them the use of thiourea (TU), chlorine, thiosulfate, bromine and iodine are being seriously investigated. In this investigation, the extraction of gold from auriferous sulfidic copper ores of Küre Region was studied by using thiourea as an alternative process to cyanidation. TU leaching results on auriferous sulfidic copper ores were presented here including the fundamental of thiourea leaching system. Thiourea leaching of gold and silver was initially discussed by Plaksın - Kozhukkova, 1941, and has attracted the interest of metallurgists since early 1960's. The literature contains several studies concerning different aspects of thiourea leaching. Previous research studies can be classified in to three groups: 1) The fundamentals of thiourea leaching based on chemistry and hydrometallurgy, 2) Laboratory and pilot plant tests to develop a thiourea process for treating ore and high grade precious metals bearing ores and concentrates, 3) Small scale industrial application studies. The extraction of gold and silver by thiourea is a viable alternative to cyanidation in specific instances. Its basic advantages over cyanidation can be summarized below: * TU leaching is carried out under acidic conditions, Au and Ag can be recovered from materials which are unstable in alkaline solutions and/or react directly with cyanide, vm* TU is compatible with small quantities of dissolved metals such copper andiron, * TU is only slightly toxic as its ultimate degradation products are elemental sulphur and cyanamid, a material which can be used as a fertiliser, and * Leaching kinetics of thiourea is approximately five times faster than cyanidation. An attempt has been made in this research to determine TU leaching parameters for auriferous sulfidic copper ores. Experimental works subjected in this thesis, were performed on a sample of 100 % to minus 0.038 mm. The ore sample used in the experimental work, which was taken from Küre Aşıköy deposit, assayed 19.47 g/t Au, 15.5 g/t Ag and 1.45 Cu. From size determination and fractional contents, it was seen that gold and copper contents slightly increases by increasing mesh of ore. The reduction of particle size caused higher dissolution rates such mat below 38 mikrons of material size, the dissolution recovery of gold and silver was 49.4 % and 22.6 % respectively. Numerous bottle roll thiourea leach tests as well as several open- beaker tests were conducted to determine the effects of the various parameters on gold and silver extraction and reagent consumption. During the tests, treated samples, in dilute sulphuric acid-ferrous sulphate solution, to remove dissolved copper from solution, and untreated samples were used. At the end of the leaching steps, the slurries were filtered, washed with water and dried leach residues were analyzed for Au and Ag, using fire assay and for copper using inductively coupled plasma spectrophotometry. Excess thiourea in the solutions was determined with a mercuric nitrate titrant and diphenly carbazide indicator. All pH-Eh measurements were done by using an Oregon, EA-940 expandable ion analyzer. Potential values were reported with respect to silver-silver chloride (4M) electrode. DCBefore thiourea leaching of the treated ore; several parameters including degradation of reagent, Eh-pH changing, influences of copper ions which would be reacted with thiourea and caused reagent consumption, and oxidant concentration were investigated. The presence of Cu+2 ions in the solution above 10 mg/1 was found to severely affect the consumption of thiourea via dissolution of the reagent. The dissolution rate of copper from the ore was found to be between 5.7 and 8.0 % under various sulfuric acid and ferrous sulfate concentrations. It was observed that while the acidity of the solution did not have a marked effect on the dissolution, an increase in the ferrous sulfate concentration reduced the dissolution of copper. Optimum copper dissolving conditions were determined by using 0.05 M sulfuric acid and 0.05 M ferrous sulfate concentration in solution having initial pH of around 1.0 and redox potential between 280 and 300 mV with eight hours of dissolving time. Copper dissolution on three stage leaching step which included dilute sulfuric acid-ferric sulfate, water and thiourea leaching was investigated. It was found that the copper dissolution during thiourea leaching is between 0.5 and 1.9 %, while 6.4-7.5 % of total copper is dissolved in the first stage. After the removal of copper ions from solutions end of the second stage, 24 hours thiourea leaching with two different concentrations (0.13M- 0.39 M) were carried out. It was seen that gold dissolution recoveries are maintained around 40-42 % in these tests while there was no significiant dissolution of silver. Degradation of acidified thiourea solution with adding of different amounts of ferric sulfate was determined. The addition of the ferric ion into the solution acidified solution gives a characteristics red color defining the complex formation of thiourea and ferric ion. On the other hand ferric ions play very important role for oxidizing of thiourea to formamidine disulfite in certain levels. In our experimental conditions, efficient ferric sulfate concentration in order to oxidizing of thiourea was determined between 7.5 and 10 g/1 values. XAfter optimizing of thiourea concentration, leaching time and influence of sodium sulfite due to the preventing of excess thiourea degradation, maximum gold dissolution with 69.1 % recovery was obtained while thiourea consumption is around 19.6 kg/t of ore. During all tests it was observed that of certain amount of thiourea remains in the solution. These remaining values of thiourea in solution are approximately half of initial thiourea amounts. In order to use the unreacted reagent, a multistage leach was setup. Three stages of fresh leach solution were introduced to the individual leach residue to determine of maximum gold dissolution, and two stages pregnant solution were introduced to fresh sample. During these tests with untreated ore sample, even though gold dissolution recovery increased up to 83.4 %, thiourea consumption also increased up to 55.4 kg/t of ore end of three stages. By combining of theoretical considerations and experimental results from this research, it was concluded that direct thiourea leaching process will be uneconomical in comparison with thiourea leaching on treated sample while the latter has numerous advantages over the former. XI